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Electrical  World         The  Engineering  andMining  Journal 
Engineering  Record  Engineering  News 

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Signal  Engineer  American  Engneer 

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Metallurgical  and  Chemical  Engineering  Power 


COAL  MINE  SURVEYING 


BY 

A.  T.  SHURICK 

ASSOCIATE    EDITOR,    COAL   AGE 


MINING  METHODS 

SELECTED  FROM 

"MINING  WITHOUT  TIMBER" 

BY 
ROBERT  BRUCE  BRINSMADE,  B.  S.,  M.  E. 


PRACTICAL  SHAFT 
SINKING 

BY 
FRANCIS  DONALDSON,  M.  E. 

CHIEF   ENGINEER,    THE    T.    A.    GILLE8PIE    COMPANY 


McGRAW-HILL  BOOK  COMPANY,  INC. 
239  WEST  39TH  STREET,  NEW  YORK 

6  BOUVERIE  STREET  , LONDON,  E.  C. 


Enginecrfnt 
Library 


2.7s 


COAL  MINE  SURVEYING 

BY 

A.  T.  SHURICK 

ASSOCIATE    EDITOR,    COAL   AGE 


COPYRIGHT,  1914,  BY 
A.  T.  SHURICK 


PREFACE 

It  has  not  been  attempted  to  make  this  book  a  "treatise"  in  any 
sense  of  the  word.  With  due  respect  to  the  numerous  comprehensive 
works  on  this  subject  it  has  been  the  experience  of  the  author  that 
too  little  consideration  is  given  to  the  problems  arising  in  the  com- 
monplace routine  of  every-day  work.  In  this  respect  it  is  believed 
the  subject  has  been  handled  in  a  distinctive  manner.  Average 
practice  has  in  every  case  been  given  preference  to  abnormal  condi- 
tions involving  problems  in  precise  surveying.  Space  has  been  freely 
devoted  to  what  may  seem  unimportant  details,  but  which  are 
none  the  less  items  that  will  do  much  to  facilitate  the  work  in  the 
mines. 

Particular  emphasis  has  been  placed  on  the  proper  care  and  ad- 
justment of  instruments  regarding  which  there  is  a  surprising  lack 
of  thorough  knowledge  among  most  engineers.  While  the  modern 
instruments  of  reputable  manufacturers  seems  to  have  attained  the 
acme  of  perfection,  the  transit  and  level  are  still  delicate  mechanisms 
susceptible  to  many  inaccuracies.  Their  relatively  high  cost,  and 
more  especially,  the  peculiarly  hard  usage  to  which  they  are  sub- 
jected in  underground  work  seems  to  justify  considerable  elabora- 
tion along  this  line.  Believing  that  the  best  information  of  this 
kind  is  obtained  at  the  point  of  manufacture,  the  instrument  makers 
have  been  quoted  exclusively  in  this  portion  of  the  work. 

In  reviewing  the  manuscript  for  this  book  as  it  is  turned  over  to 
the  printer,  the  author  feels  that  he  has  little  claim  to  authorship. 
Opinions  and  methods  described  in  the  various  technical  journals 
and  other  publications  have  been  freely  quoted  except  where  of  local 
interest  only.  But  in  all  such  cases  a  judicious  selection  of  accepted 
authorities  only  have  been  used,  and  the  author's  prerogative  of 
wielding  the  blue  pencil  has  been  freely  exercised. 

A.  T.  SHURICK. 

NEW  YORK,  March,  1914. 


CONTENTS 


PAGE 
PREFACE  .  .    vii 


CHAPTER  I 

PRINCIPLES   OF  SURVEYING 

AZIMUTHS,  BEARINGS  AND  COURSES 'i 

Azimuths — Bearings  and  Courses. 

LATITUDES  AND  DEPARTURES 2 

LEVELING 8 

Vertical  Angle  Method— Direct  Method  by  the  Level. 

CHAPTER  II 

SURVEYING  INSTRUMENTS  AND  ACCESSORIES 

THE  TRANSIT 10 

The  Telescope — The  Cross  Wires — The  Vernier  Graduations  on  the 

Horizontal  Circle. 
THE  LEVEL ' 18 

The  Telescope — Level  Bar  and  Telescope  Wye — The  Tangent  Clamp 

Screw — The  Leveling  Head. 

LEVEL  ROD 20 

STEEL  TAPES 23 

PLUMB  BOBS ' 24 

CHAPTER   III 

CARE  OF  INSTRUMENTS 

-IN  GENERAL 26 

CARE  OF  CENTERS  AND  GRADUATIONS ;    .    .   .    .  29 

TELESCOPE  LENS 30 

LUBRICATING 31 

LEVEL  BUBBLES 33 

Mounting  Spirit  Levels. 

REPLACING  CROSS  WIRES 35 

ACCIDENTS  AND  PRECAUTIONS  IN  THE  USE  OF  INSTRUMENTS 36 

TRANSPORTATION  OF  INSTRUMENTS 39 

ix 


X  CONTENTS 

CHAPTER  IV 

ADJUSTMENT  OF  INSTRUMENTS 

PAGE 

THE  TRANSIT 42 

The  Bubbles— To  Make  the  Adjustment  for  Parallax — To  Make 
the  Vertical  Cross  Wire  Perpendicular  to  the  Plane  of  the  Horizontal 
Axis— To  Adjust  the  Vertical  Wire — To  Determine  Whether  the  Stand- 
ards are  the  Same  Height — To  Adjust  the  Level  to  the  Line  of  Colli- 
mation  of  the  Horizontal  Wire. 

THE  LEVEL 45 

To  Make  the  Adjustment  of  the  Horizontal  Wire— To  Center  the 
Eyepiece — To  Adjust  the  Spirit  Level  to  the  Telescope — To  Make  the 
Lateral  Adjustment  of  the  Spirit  Level — To  Make  the  Adjustment  of  the 
Level  Bar — To  Adjust  the  Horizontal  Wire  so  that  the  Line  of  Sight 
will  be  Parallel  to  the  Spirit  Level. 

CHAPTER  V 

ORGANIZING  AND  EQUIPPING  THE  FIELD  PARTY 
The  Chief  of  the  Party — The  Transitman — Chainman  and  Backsight.     48 

CHAPTER  VI 
ENTRY  SURVEYING 

Picking  up  the  Starting  Stations — Setting  Up  the  Transit — Turning 
Angles— Straight  Line  Work— Chaining— Taking  Side  Notes— Making 
Checks — Kinds  of  Stations — Painting  Stations — Numbering  Stations.  52 

CHAPTER   VII 
KEEPING  SURVEY  NOTES 

TRANSIT  NOTES 65 

LEVEL  NOTES .      72 

CHAPTER   VIII 

SOME  PROBLEMS  IN  SURVEYING 
Selected  Problems  Commonly  Met  in  Coal  Mine  Surveying  ....     76 


COAL  MINE  SURVEYING 

CHAPTER  I 

PRINCIPLES  OF  SURVEYING 

The  art  of  surveying  consists  essentially  in  the  relative  location  of 
different  points  with  respect  to  each  other.  The  average  mining 
man  is  too  prone  to  look  upon  the  work  of  the  engineer  (or  surveyor 
as  he  is  perhaps  more  commonly  termed  in  the  mining  regions)  as 
something  uncanny  and  bordering  on  the  supernatural.  He  is 
usually  regarded  as  an  expensive  luxury,  or  a  necessary  evil.  No 
tangible  evidence  of  his  labors  is  notable  on  the  tonnage  sheets 
of  the  mine  he  is  working  in — in  fact,  it  is  more  often  the  case  that 
he  leaves  a  trail  of  profane  "skinners"  who  have  been  delayed,  in 
his  wake.  Nevertheless  he  has  come  to  stay  and  each  year  finds  him 
occupying  a  stronger  foothold,  until  now  one  of  the  best  criterions 
of  an  efficient  management  is  shown  in  the  excellence  of  its  en- 
gineering practice. 

AZIMUTHS,  BEARINGS  AND  COURSES 

To  determine  the  location  of  a  point  it  is  obvious  that  two  things 
are  required,  the  direction  and  distance  from  an  already  known 
point.  Directions  are  designated  as  azimuths,  bearings  and  courses. 
For  convenience  and  in  order  to  have  a  worldwide  standard,  so  that 
all  surveys  of  whatever  character  may  be  readily  adjusted  to  each 
other,  the  true  north  and  south  meridian  has  been  adopted  as  the 
basis  from  which  all  computations  referring  to  direction  are  made. 

Azimuths. — In  Fig.  i,  it  will  be  noted  that  azimuths  cover  the 
entire  range  of  the  circle  from  o°  to  360°.  It  is  the  angle  which  the 
line  makes  with  the  true  meridian,  measured  to  the  right  in  the  same 
direction  as  the  hands  of  a  watch.  Thus  as  will  be  noted  in  the 
figure,  line  C  lies  98°  18'  to  the  right  of  the  meridian;  line  B,  205°  30' 
and  line  A  338°  21'.  It  is  important  that  the  student  thoroughly 
master  the  principle  of  azimuths  for  the  reason  that  they  greatly 
simplify  certain  processes  in  surveying  which  are  to  be  described 
later. 

Bearings  and  Courses. — These  are  terms  that  are  used  indis- 


2  COAL  MINE  SURVEYING 

criminately  to  designate  the  direction  of  a  line  in  one  of  the  four 
quadrants  of  the  circle,  as,  northeast,  northwest,  southeast,  and 
southwest.  Sometimes  the  expression  also  implies  distance  as 
well  as  direction,  as  for  instance  the  bearing  276ft.  N.  46°  i8'E.  Bear- 
ings differ  from  azimuths  in  that  they  are  measured  both  to  the 
east  and  west  of  the  north  and  south  meridian  and  that  they  never 
exceed  90°  in  value.  Thus  in  Fig.  i  the  bearing  of  C  equals  the 
angle  SOC,  or  the  difference  between  the  azimuth  of  SO  (180°)  and 
OC  (98°  18')  which  equals  81°  42';  and  since  the  line  lies  between 


PIG.    I.      SKETCH  SHOWING  VARIOUS   AZIMUTHS   AND  BEAKINGS. 


the  due  east  and  due  south  meridians  the  bearing  is  obviously  S. 
81°  42'  E.  The  bearings  of  the  other  lines  in  the  figure  are  arrived 
at  by  the  same  process. 

LATITUDES  AND  DEPARTURES 

When  the  survey  of  a  mine  has  been  completed  the  problem 
then  arises  of  making  an  exact  and  accurate  reproduction  of  the 
mine  workings  on  a  small  scale.  In  other  words,  building  the  map. 
There  are  two  general  methods  of  accomplishing  this.  By  plotting 
with  a  protractor  and  by  computing  the  latitudes  and  departures 
and  plotting  the  survey  from  them. 


PRINCIPLES  OF  SURVEYING  3 

Plotting  by  means  of  a  protractor  was  at  one  time  a  popular 
method  of  working  but  it  is  now  obsolete  and  has  been  practically 
abandoned.  The  possibility  for  error  is  too  great  by  this  method 
unless  great  care  is  exercised  and  even  then  it  is  not  to  be  relied  upon 
for  any  very  extended  work.  The  method  is  still  used  in  laying 
out  short  or  approximate  surveys  but  it  is  rather  dangerous  for  the 
reason  that  any  mistakes  in  plotting  one  chord  are  carried  through 
the  balance  of  the  work,  and  are  accumulative. 

In  latitudes  and  departures  the  northing  or  southing  and  the 
easting  or  westing  of  each  bearing  is  computed  by  means  of  sines  and 
cosines;  that  is  the  distance  due  north  or  south  and  due  east  or  west 
(depending  upon  which  direction  the  bearing  is  in).  The  algebraic 
sums  of  all  the  bearings  in  the  survey  are  then  obtained  and  each 
station  plotted  according  to  its  distance  due  north  or  south  and  due 
east  or  west  of  an  assumed  zero  point,  termed  the  zero  of  coordinates. 
The  method  is  probably  best  described  by  means  of  an  example. 

Referring  to  the  accompanying  Table  I  it  will  be  assumed  that  the 
field  notes  for  a  survey  as  given  in  the  columns  under  "Sta.," 
"Azimuth"  and  "Distance"  have  been  turned  into  the  office  and  it 
is  desired  to  traverse  them  and  find  the  latitude  and  departure  of 
each  station.  It  is  first  necessary  to  reduce  the  azimuths  to  bearings, 
this  having  been  already  explained. 

The  latitude  and  departure  of  each  course  is  then  found  by 
multiplying  the  length  of  the  course  by  the  cosine  and  sine,  re- 
spectively. Thus,  the  latitude  of  the  first  course  being  a  northing 
(N)  and  the  departure  an  easting  (£),  they  are  found  as  follows: 


N=i45  cos  43°  i8'  =  i45Xo.72777  =  io5.53  ft. 
£=145  sin  43°  18'  =  145X0.68582  =  99.44  ft. 

In  this  manner,  the  northing  or  southing  and  the  easting  or  westing 
of  each  course  is  calculated  and  written  in  the  proper  column  under 
"Singles"  as  shown  in  the  table. 

Having  found  all  the  single  latitudes  and  departures  for  all  the 
bearings  we  next  proceed  to  obtain  the  "Doubles."  The  single 
latitudes  and  departures  are  the  distances  which  that  particular  course 
goes  north  or  south  and  east  or  west,  while  the  doubles  are  the  total 
distances  north  or  south  and  east  or  west  of  each  point  from  the 
zero  of  coordinates.  The  doubles  are  obtained  by  simply  adding 
or  subtracting  the  singles,  as  the  case  may  be. 

Referring  to  Fig.  2  which  is  a  plot  of  the  survey  under  considera- 
tion, the  lines  NS  and  WE  are  due  north  and  south,  and  east  and  west, 


COAL  MINE  SURVEYING 


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PRINCIPLES  OF  SURVEYING  5 

respectively.  Their  intersection  is  the  assumed  zero  of  coordinates 
and  is  therefore  the  point  that  has  neither  a  northing,  southing,  east- 
ing or  westing.  Sta.  o,  which  is  the  starting  point  of  the  survey,  is 
known  from  a  previous  survey  to  have  a  northing  of  108.00  ft.  and 
a  westing  of  138.00  ft.  as  will  be  observed  by  reference  to  the  map. 
Accordingly  we  set  down  108.00  in  the  northing  column  under  the 
doubles  and  138.00  in  the  westing  column. 

From  Sta.  o  to  i,  according  to  the  single  latitude  and  departure  for 
this  course,  we  go  105.53  ft.  north  and  since  the  double  latitude  is 


FIG.    2.      SKETCH  OF  A  SURVEY  PLOTTED  BY  LATITUDES  AND  DEPARTURES. 


already  north  this  is  added  to  it,  giving  a  latitude  of  213.53  ft- 
be  noted  on  both  the  table  and  map.  Similarly,  the  course  o  to  i 
takes  us  99.44  ft.  east.  In  this  case,  however,  our  double  is  a  westing 
so  that  we  subtract  instead  of  adding  as  in  the  previous  instance. 
Thus  we  have  the  rule:  When  the  single  latitude  or  departure  is  in 
the  same  direction  as  the  double  it  is  always  added  and  when 
different  it  is  always  subtracted. 
On  the  plot,  Fig.  2,  all  of  the  latitudes  and  departures  of  the 


6  COAL  MINE  SURVEYING 

different  points  have  been  noted.  Beginning  at  Sta.  i  and  following 
around  to  Sta.  4  it  will  be  noted  that  the  stations  lie  to  the  north  of 
the  east  and  west  line  and  hence  are  northings  as  will  be  observed  by 
reference  to  the  double  latitude  and  departures.  Between  Stas.  4 
and  5  the  traverse  crosses  to  the  south  and  continues  there  up  to  Sta. 
9  as  will  be  seen  by  reference  to  both  the  map  and  the  table.  Fol- 
lowing the  departures  around  in  a  similar  manner  we  find  that  the 
survey  crosses  from  a  westing  to  an  easting  between  Stas.  i  and  2, 
shows  another  small  westing  at  Sta.  3  and  then  continues  an  easting 
up  to  Sta.  7;  between  Stas.  7  and  8  it  changes  to  a  westing  and 
remains  so  the  remainder  of  the  survey.  It  will  also  be  noted  that 
this  is  a  tie  survey;  that  is  it  completes  a  circuit  and  ends  at  the 
point  of  beginning  and  therefore  must  balance.  Referring  to  the 
double  latitudes  and  departures  we  find  that  the  final  coordinates 
for  the  Sta.  o  correspond  exactly  with  those  at  the  beginning,  and 
therefore  a  perfect  check  has  been  obtained.  It  might  be  well  to 
add  that  a  survey  in  practice  which  checked  perfectly  would  be  open 
to  suspicion  as  such  a  tie  is  seldom  or  never  obtained. 

The  following  is  a  description  of  the  Consolidation  Coal  Go's, 
methods  of  computing  latitudes  and  departures: 

The  first  step  in  the  office  work  is  to  reduce  all  slope  measure- 
ments to  horizontals.  These  calculations  are  performed  in  dupli- 
cate, one  set  by  a  table  of  Gurden  and  Naturals  and  the  other  by  log- 
arithms and  the  results  check  one  against  the  other.  The  trans- 
itman  in  the  meantime  transfers  his  data  as  far  as  possible  to  tra- 
verse sheets,  copy  of  which  is  shown  herewith.  Usually  each  mine 
entry  has  an  individual  traverse  sheet  for  the  stations  it  contains. 
After  the  horizontals  are  copied  into  the  notebook  and  checked, 
the  courses  and  corresponding  distances  are  then  copied  into  a 
separate  set  of  calculation  books  and  the  latitude  and  departure 
differences  worked  out  by  the  same  method  employed  for  horizontals, 
and  also  checked.  The  results  are  copied  on  the  traverse  sheets  and 
the  total  latitude  and  departures  worked  out  and  checked  before 
plotting. 

No  survey  is  permitted  to  stand  until  approved  by  the  division 
engineer  in  charge.  His  approval  is  shown  on  the  space  provided 
for  his  signature  at  the  bottom  of  the  traverse  sheet.  All  sheets  are 
carefully  referenced  from  one  to  another  when  the  surveys  have  any 
connections  whatever.  A  separate  folio  or  binder  is  used  to  hold  the 
sheets  for  each  mine  and  they  are  numbered  consecutively  from 
"one  up";  the  sheets  are  of  course  all  indexed. 


PRINCIPLES  OF  SURVEYING 


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8  COAL  MINE  SURVEYING 

LEVELING 

In  addition  to  the  location  of  points  in  a  horizontal  plane  there  is 
also  the  necessity  of  determining  the  relative  elevation  of  different 
portions  of  the  mine  workings.  This  branch  of  surveying  has  not 
received  so  much  attention  as  the  other  and  in  fact  is  often  entirely 
neglected  unless  some  special  occasion  arises  where  it  is  required. 
It  is  none  the  less  important  however  and  will  no  doubt  find  a  more 
general  application  in  time. 

There  are  two  distinct  methods  of  leveling — indirectly  by  vertical 
angles  and  directly  by  the  level,  an  instrument  designed  especially 
for  this  work.  The  vertical  angle  method  is  not  in  much  favor  on 
outside  work  and  except  in  special  cases  it  is  seldom  or  never  used. 
But  in  the  narrow  confines  of  a  heavily  pitching  seam  of  coal  the 
adoption  of  this  system  becomes  imperative  so  that  the  coal  engineer 
must  be  prepared  to  handle  it. 

Vertical  Angle  Method. — In  this  system  the  difference  in  elevation 
between  any  two  points  is  arrived  at  by  measuring  the  vertical  angle 
between  them  by  means  of  an  ordinary  transit,  and  the  slope  distance. 
We  then  have  the  hypotenuse  of  a  right  angled  triangle  and  one  angle 
given  so  that  the  vertical  distance  is  easily  computed  by  multiplying 
the  sine  of  the  angle  by  the  distance.  It  should  be  remembered 
in  this  connection  that  the  average  transit  is  not  usually  equipped 
with  a  sufficiently  accurate  vertical  arc  to  insure  absolute  correct 
results  on  work  of  this  kind.  In  fact  the  arc  should  be  as  large  and 
rigid  as  in  the  case  of  the  horizontal  plates  where  equally  accurate 
results  are  desired.  In  ordering  an  instrument  it  is  well  to  consider 
the  geological  features  of  the  district  carefully  and  where  the  measures 
are  steeply  inclined  a  special  vertical  arc  should  be  specified. 

Direct  Method  by  the  Level.— As  already  noted  this  is  the  system 
in  most  general  use  and  the  one  to  be  given  the  preference  where 
accuracy  is  desired  and  the  inclinations  not  too  abrupt.  It  consists 
essentially  in  carrying  forward  a  series  of  horizontal  planes,  either 
up  or  down  as  the  case  may  be,  as  shown  in  Fig.  4. 

When  properly  adjusted  the  level  is  an  instrument  that  may  be 
turned  in  any  direction  in  a  horizontal  plane  and  will  always  show 
identically  the  same  elevation.  Thus  in  Fig.  4  let  it  be  assumed 
that  it  is  desired  to  ascertain  the  difference  in  elevation  between 
the  points  A  and  £  on  a  hillside.  The  instrument  is  set  up  at 
any  point  B  not  too  high  so  that  the  rod  will  not  be  seen,  and  a 
reading  on  the  latter  taken  at  A.  Say  this  reading  shows  11.58  ft.; 


PRINCIPLES  OF  SURVEYING  9 

it  is  clear  then  that  the  instrument  is   this  distance  above  the 
point. 

The  rod  is  then  moved  ahead  to  another  point  at  about  the  same 
elevation  as  the  instrument  and  another  reading  taken.  Assuming 
the  reading  at  this  point  to  be  0.42  ft.  this  indicates  that  the  new 
point  C  is  this  distance  below  the  instrument  and  the  difference 
between  11.58  ft.  and  0.42  ft.  or  11.16  ft.  above  the  starting  point  A. 
Moving  the  instrument  up  to  D  this  time,  the  operation  is  repeated 


FIG.    4.      SKETCH   SHOWING   LEVELING   PROCESS. 

as  follows:  Assuming  the  rod  reading  on  C  from  the  new  set-up 
at  D  shows  10.50  the  height  of  the  instrument  above  the  initial  point 
is  then  11.16  plus  10.50  or  21.66  ft.  Finally  placing  the  rod  on  the 
desired  point  E  we  find  this  to  be  say  0.12  ft.  below  the  instrument 
so  that  the  difference  in  elevation  between  A  and  E  is  21.66  minus 
0.12  or  21.54  ft.  And  so  the  operation  maybe  extended  indefinitely. 


CHAPTER  II 
SURVEYING  INSTRUMENTS  AND  ACCESSORIES 

In  the  ordinary  surveying,  as  practised  in  the  coal  mines  of  this 
country,  comparatively  simple  equipment  suffices;  in  fact  it  might 
be  said  the  simpler,  the  better.  The  principal  instruments  used 
are  the  transit  and  level,  together  with  such  accessories  as  leveling 
rods,  sight  rods,  plumb  bobs,  steel  tapes,  etc. 

The  average  layman  invariably  has  the  compass  also  associated 
with  anything  in  connection  with  surveying.  As  a  matter  of  fact, 
the  compass  has  been  practically  abandoned  altogether  for  use  in 
this  connection,  and  could  be  dispensed  with  entirely  with  little 
inconvenience.  However,  it  is  of  some  use  as  an  approximate  check  on 
the  work  with  the  vernier,  and  is  still  a  part  of  nearly  all  transits 
probably  more  because  it  has  become  a  custom  and  also  because  the 
space  so  utilized  could  not  be  applied  to  an  advantage  in  any  other 
way. 

THE  TRANSIT 

The  most  important,  and  at  the  same  time  most  complicated 
and  expensive  instrument  used  on  the  mine  survey  is  the  transit. 
Illustrations  of  two' popular  makes  of  transits  are  shown  herewith. 
The  first,  Fig.  5,  is  a  halftone  of  the  Kueffel  &  Esser  instrument,  while 
the  second,  Fig.  6,  is  a  cross  section  through  the  center  of  a  C.  L. 
Berger  transit.  In  both  illustrations  all  the  parts  of  the  two  instru- 
ments are  numbered,  and  the  names  of  these  will  be  found  in  accom- 
panying tables. 

The  transit  is  an  instrument  designed  for  measuring  angles  in 
both  horizontal  and  vertical  planes,  although  its  use  is  more  com- 
monly confined  entirely  to  the  former.  To  adequately  fulfil  its 
purpose,  the  instrument  must  be  rigidly  constructed,  with  all  parts 
in  absolute  adjustment.  Rapid  advances  in  the  method  of  manu- 
facturing instruments  have  been  made  in  the  past  decade,  with  the 
result  that  the  modern  transit  is  a  model  of  accuracy  and  conven- 
ience and  will  successfully  withstand  as  much  ill  usage  as  may  rea- 
sonably be  expected  from  so  delicate  a  machine. 

Beginning  at  the  bottom  and  working  up,  it  will  be  noted  on  refer- 


SURVEYING  INSTRUMENTS  AND  ACCESSORIES 


ring  to  the  drawing,  Fig.  6,  that  the  tripod  head  is  shown  at  i  with 
various  parts  of  the  tripod  at  2,  3,  4  and  5.  The  tripod  is  entirely 
distinct  from  the  instrument,  but  the  head,  i,  is  equipped  with  screw 
threads,  as  shown  in  the  drawing,  on  to  which  the  instrument  foot 
plate  screws.  The  foot  plate,  or  tripod  plate,  is  shown  at  6  in  Fig. 
6,  and  13  in  Fig.  5. 

1.  Objective  head 

2.  Telescope 

3.  Telescope    axis 

end  cap 

4.  Vertical  circle 

5.  Vertical     circle 

vernier. 

6.  Standard 

7.  Plate  level 

8.  Horizontal  limb 

9.  Lower  clamp  tan- 

gent screw 

10.  Lower  clamp 

11.  Leveling  head 

12.  Leveling  screws 

13.  Tripod  plate 

14.  Pinion  head  (ob- 

ject      focusing 
screw) 

15.  Standard  cap 

16.  Telescope  clamp 

17.  Reticule    adjust- 

ing screw 

18.  Eyepiece     focus- 

ing screw 

19.  Eyepiece     focus- 

ing lock  nut 

20.  Eye  end  ring 

21.  Eyepiece  cap 

22.  Telescope     level 

support 

23.  Telescope      level 

adjusting  screw 

24.  Telescope  level 
23.  Telescope  clamp 

tangent  screw 

26.  Compass     cover 

glass  ' 

27.  Capston      head 

pinion 

28.  Vernier     cover 

glass 

29.  Upper  clamp  tan- 


13 


gent  screw 

30.  Upper  clamp 

31.  U  pper    clamp 

collar 

32.  Clamp  collar 

33.  Lower    clamp 

screw 

34.  Half  ball  joint 

35.  Shifting  plate 

36.  Leveling    screw 

shoe 


FIG.    5.      TYPICAL  TRANSIT. 


The  shifting  plate  of  the  instrument  in  Fig.  6  is  shown  at  n. 
The  purpose  of  this  is  to  permit  of  the  final  movement  necessary  to 
get  the  instrument  precisely  over  or  under  the  point,  this  being  ac- 
complished by  loosening  the  level  screws  9,  so  that  the  plate  hangs 
relatively  loose.  To  avoid  indentations  in  the  foot  plate,  by  the 
leveling  screws  turning  directly  upon  the  plate,  leveling  screw 


I2  COM,  MINE  SURVEYING 

cups  are  provided,  as  shown  at  10.     Turning  to  Fig.  5,  the  leveling 
screw,  and  cups  are  shown  at  12  and  36,  respectively. 

The  plumb  bob  suspending  cup  with  the  accompanying  chain  and 
hook  are  shown  at  12  and  13,  respectively,  in  Fig.  6;  these  are 
arranged  so  that  a  bob  hung  in  the  hook  13  will  hang  precisely  under 


Tripod  head 

Tripod  bolt 

Tripod  bolt  nut 

Tripod  bolt  washer 

Tripod  leg 

Foot  plate 

Ball-and-socket  joint 

Leveling  head 

Leveling  screws 

Leveling  screw  cups 

Shifting  plate 

Plumb  t>ob  suspending  cup  ; 

Plumb  bob  chain  and  hook 

Repeating  center 

Clamp  collar 

Lower  clamp 

Lower  clamp  thumb  screw 

Lower  clamp  tangent  screw 

Lower  clamp  spring,  not  shown 

Lower  clamp  piston,  not  shown 

Lower  clamp  cap,  not  shown 

Inner  center 

Horizontal  circle 

Vernier  plate  clamp 

Vernier  plate  clamp  screw 

Vernier  plate 

Vernier  plate  tangent  bracket 

Vernier  plate  tangent  screw 

Vernier  plate  tangent  spring 

Tangent  spring  piston 

Tangent  spring  cap 

Horizontal  vernier 

Veriner  cover  glass 

Vernier  shade  frame 

Vernier  shade  glass 

Compass  cover  glass 

Needle 

Needle  pivot 

Needle  lifter 

Needle  lifter  screw  head 

Front  plate  level 

Front  plate  level  vial 

Front  plate  level  adjusting  n 

Front  plate  level  rocker 

Standard  with  adjusting  screw 

Standard  adjustable  wye 

bearing 

Standard  cap  and  screws 

Side  level 

Side  level  adjusting  screw 

Sjde  level  fastening  screw 

Side  level  rocker 

Side  level  vial 

Telescope  axis 


-71  SI: 


54-  Telescope  barrel 

55.  Eyepiece  mounting 

56.  Spiral  groove  screw 

57.  Terrestrial  eyepiece 

58.  Eyepiece  cap 

59.  Eyepiece  ring 

60.  Wire  recticule 

61.  Wire  recticule  adjusting 
screws 

62.  Pinion  and  washer,  not 
shown 

63.  Pinion  head 

64.  Pinion  head  screw 

65.  Pinion  head  saddle 

66.  Pinion  head  saddle  screws 

-  67.  Object  slide 

*  68.  Object  head 

47  69.  Object  glass  cell 
-46 7°'  Object  glass 

71.  Object  slide  dust  guard 

72.  Sun  shade 

L     73.  Telescope  clamp 
'  74.  Telescope  clamp  screw 
75.  Telescope  clamp  washer 


Telesco 


pe  t 
l  cir 


tangent 

77.  Vertical  circle 

78.  Vertical  circle  guard 

79.  Vertical  circle  guard  ser 

80.  Vertical  vernier 

81.  Vertical  vernier  screws 


FIG.   6.      CROSS   SECTION   OF   A   TRANSIT. 


the  theoretical  center  line  of  the  instrument.  The  ball-and-socket 
joint,  which  is  something  on  the  order  of  a  flexible  coupling,  is  shown 
at  7  in  Fig.  6,  and  34  in  Fig.  5.  This  allows  the  upper  part  of  the 
instrument  to  assume  any  necessary  angle  of  inclination  to  the 
tripod  plates,  so  that  it  will  be  absolutely  level. 
The  leveling  head,  through  which  the  leveling  screws  work,  is 


SURVEYING  INSTRUMENTS  AND  ACCESSORIES  13 

shown  at  ii  in  Fig.  5,  and  8  in  Fig.  6.  In  Fig.  5  the  lower  clamp  for 
clamping  the  horizontal  plate  is  shown  at  33,  and  the  tangent  screw 
at  9,  these  same  devices  being  shown  at  17  and  18,  respectively  in 
Fig.  6.  The  upper  or  venier  plate  clamp  is  shown  at  30,  with  the 
tangent  screw  at  29,  in  Fig.  5,  the  same  thing  being  shown  at  25  and 
28  in  the  separate  detail  in  Fig.  6. 

This  latter  detail  also  shows  the  principle  upon  which  the  slow 
motion  screws  work.  It  will  be  noticed  that  this  is  essentially  a 
small  cylinder  containing  a  spring  29,  held  in  place  at  one  end  by 
the  milled  head  cap  31,  and  at  the  other  by  the  tangent  spring 
piston  30.  The  tendency  is  naturally  to  push  the  piston  out  so  that 
the  latter  is  constantly  bearing  against  the  clamp  upon  which  it 
acts.  Thus  by  turning  the  tangent  screw  28,  the  cylinder  is  either 
forced  in  or  out  as  the  case  may  be.  By  this  means,  an  infinitesimal 
movement  of  the  vernier  plate  can  be  obtained  so  small  that  it  cannot 
be  detected  by  the  naked  eye. 

At  23  in  Fig.  6  is  shown  the  horizontal  plate  or  limb  which  is 
also  shown  in  Fig.  5  at  8.  The  vernier  plate  is  shown  at  26  in  Fig.  6, 
with  the  vernier  at  32,  and  the  vernier  glass  at  33,  this  latter  being 
shown  in  Fig.  5  at  28.  The  vernier  shade  is  shown  at  34  in  Fig. 
6,  this  being  removed  on  Fig.  5. 

The  compass  needle  is  shown  in  Fig.  6  at  37,  and  the  lifter,  which 
brings  the  needle  tight  against  the  compass  glass  when  not  in  use, 
so  as  to  save  it  from  unnecessary  wear,  is  shown  at  39,  the  screw  head 
by  which  this  is  accomplished  being  shown  at  40,  and  the  glass  at 
36,  this  latter  also  being  shown  in  Fig.  5  at  26. 

All  transits  are  equipped  with  two  level  bubbles,  placed  at  right 
angles  to  each  other  so  that  the  instrument  can  be  leveled  in  both 
directions.  In  Fig.  6,  these  levels  are  shown  at  41  and  48,  and  their 
relative  positions  may  be  more  adequately  noted  in  Fig.  5,  one  being 
shown  at  7,  and  the  other  about  halfway  up  the  standard  6.  The 
standard  in  Fig.  6  is  shown  at  45,  and  the  adjustable  wye  bearings 
for  correcting  the  bearing  adjusm  en  tare  shown  at  46,  with  the  cap  at 
47;  this  latter  is  also  shown  in  Fig.  5  at  15. 

In  Fig.  5  the  vertical  circle  for  measuring  vertical  angles  is  shown  at 
4,  with  the  vernier  at  5;  this  is  also  shown  in  section  in  Fig.  6,  the 
vertical  circle  itself  at  77",  the  surrounding  guard  for  protecting 
the  circle,  which  will  also  be  observed  in  Fig.  5,  at  78,  and  the  vernier 
with  its  adjusting  screws  at  80  and  81,  respectively.  The  vertical 
circle  guard  is  held  in  place  by  milled  head  screws,  one  of  which  is 
shown  at  79. 


1 4  COAL  MINE  SURVEYING 

The  telescope  clamp  for  clamping  the  telescope  in  any  vertical 
position  desired  is  shown  at  1 6  in  Fig.  5,  and  the  tangent  or  slow  mo- 
tion screw  for  setting  same  accurately  in  place  at  25.  This  arrange- 
ment is  also  shown  in  Fig.  6,  the  clamp  at  73,  with  the  clamp  screw 
at  74,  and  the  tangent  screw  at  76. 

The  telescope  is  shown  at  2  in  Fig.  5,  the  objective  head  at  i, 
and  the  eyepiece  cap  at  21.  The  eyepiece  focusing  screw  is  shown 
at  18,  the  lock  nut  for  same  at  19,  and  the  eye  end  ring  at  20.  The 
reticule  screws  for  adjusting  the  cross  wires  are  shown  at  17.  Under- 
neath the  telescope  and  attached  to  it  is  the  telescope  level  24  which 
must  always  be  leveled  up  with  the  other  bubbles  when  setting  the 
instrument  up  under  a  station.  The  telescope  level  support  and 
adjusting  screws  are  shown  at  22  and  23,  respectively. 

The  telescope  as  shown  in  Fig.  6  gives  greater  details.  The  eye- 
piece cap  is  shown  at  58,  the  mounting  for  the  eyepiece  at  55,  and 
the  spiral,  groove  screw  at  56.  The  reticule  for  carrying  the  cross 


FIG.    7.      TYPES  OF  "CROSS  WIRES  USED  IN  LEVELS  AND  TRANSITS. 

wires  is  shown  at  60,  and  the  screws  for  adjusting  these  to  their  proper 
position  at  61.  The  barrel  of  the  telescope  is  shown  at  54,  with 
the  object  slide  just  inside  of  it  at  67.  The  pinion  head  for  adjust- 
ing the  telescope  for  any  length  of  sight  is  shown  at  63,  with  the 
pinion  saddle  at  65.  The  slide  dust  guard  for  preventing  foreign 
matter  from  getting  inside  the  telescope  is  shown  at  71,  and  the 
object  head  at  68,  with  the  object  glass  at  70,  and  the  barrel  for 
holding  same  at  69.  The  sun  shade  for  preventing  the  sun  from 
shining  directly  on  the  object  glass  when  taking  a  sight  is  shown 
at  72. 

Cross  Wires.— Some  of  the  different  styles  of  cross  wires  used  in 
the  present-day  instruments  are  shown  in  the  accompanying  illus- 
tration, Fig.  7.  The  first  style  on  the  left  is  an  unusual  type  of  cross 
hair  used  m  wye  levels,  the  idea  of  the  two  vertical  wires  being  to 
check  up  the  rodman  in  holding  his  rod  plumb.  The  second  illus- 


SURVEYING  INSTRUMENTS  AND  ACCESSORIES  15 

tration  from  the  left  represents  the  simplest  form  of  cross  wire  used, 
and  one  that  is  also  quite  popular;  it  consists  of  only  a  single  vertical 
and  horizontal  wire.  Next  to  this  is  the  stadia  wire,  which  is  iden- 
tical with  the  one  just  described,  except  that  two  additional  hori- 


FIG.    5.      TYPES    OF  VERNIERS. 

zontal  wires  have  been  added  for  making  stadia  measurements. 
The  last  figure  on  the  right  shows  the  stadia  wires  as  just  described 
with  diagonal  wires  added,  the  idea  of  the  latter  being  to  help  locate 
either  the  vertical  or  horizontal  wire  when  working  in  the  mine. 
This  is  a  valuable  addition  in  this  respect,  and  one  that  might  be  used 
to  advantage  on  all  mine  transits. 


1 6  COAL  MINE  SURVEYING 

The  Vernier. — The  vernier  is  a  device  for  accurately  measuring 
fractions  of  subdivisions  on  any  kind  of  a  scale.  Verniers  on  transits 
designed  for  reading  to  i'  are  shown  in  the  accompanying  illustra- 
tion, Fig.  8.  This  is  the  typical  style  used  on  ordinary  mine  surveys; 
in  fact,  it  might  be  said  that  it  is  used  almost  to  the  exclusion  of  any 
other  graduation. 

These  verniers  depend  upon  the  principle  that  if  29  subdivisions 
on  the  outer  circle  equal  30  subdivisions  on  the  inner  circle,  the 
difference  in  length  between  a  subdivision  on  each  scale  is  equal  to 
-fa  of  the  outer  subdivision;  since  the  outer  subdivisions  in  this 
case  are  equal  to  ?°  or  30'  this  value  is  therefore  i'. 

With  this  principle  once  clear  in  the  mind,  reading  the  vernier 
becomes  a  comparatively  simple  problem.  Thus  referring  to  the 
vernier  B  in  Fig.  8,  and  assuming  that  it  is  desired  to  know  the 
reading  on  the  outer  circle  of  figures  we  first  select  the  closest  even 
degree,  which  in  this  case  is  obviously  332°,  the  smallest  divi- 
sions on  the  outer  circle  being  2°  and  the  next  larger  full  degrees. 
It  is  also  clear  that  the  zero  on  the  vernier  (which  is  the  inside  scale) 
is  something  more  than  a  |°  or  30'  greater  than  33  2°,  so  that  we  have 
this  additional  quantity  and  all  that  remains  to  be  obtained  is  the  odd 
minutes.  To  obtain  these,  we  look  along  the  graduations  on  the 
vernier  itself,  in  the  same  direction  in  which  the  reading  is  being 
made,  that  is,  to  the  left  in  this  case,  until  we  find  a  line  that  ex- 
actly coincides  with  a  line  on  the  outer  circle,  which  in  this  case  is 
the  fourth  subdivision.  This  indicates  that  the  zero  on  the  ver- 
nier is  -gV  of  the  distance  between  the  two  subdivisions  on  the 
outer  circle,  which  as  was  already  explained  equals  4'. 

Referring  to  vernier  A,  we  find  that  this  reads  to  even  degrees, 
the  zero  on  the  outer  circle  exactly  corresponding  with  the  zero  on 
the  vernier.  Vernier  C  is  the  same,  except  that  this  reads  exactly  5°. 
In  vernier  D,  however,  we  again  find  a  reading  to  odd  minutes. 
Assuming  that  it  is  again  desired  to  read  the  outer  circle  of  figures  it 
is  readily  noted  that  the  reading  is  somewhat  greater  than  152!° 
or  152°  30'.  Glancing  along  the  inner  circle  from  the  zero  to 
the  left  we  find  the  first  line  to  correspond  at  5,  which  is  of  course, 
as^already  explained,  equal  to  5',  and  the  reading  is  therefore  152° 
35'.  In  the  same  way,  we  find  the  reading  of  vernier  E  to  be  342° 
35'.  The  verniers  here  given  are  typical  of  those  used  on  the 
average  instruments. 

Graduations  on  the  Horizontal  Circle.— The  accompanying  illus- 
tration, Fig.  9,  shows  two  types  of  graduation  commonly  used  on  the 


SURVEYING  INSTRUMENTS  AND  ACCESSORIES  17 

mine  transits.  The  outer  one  is  graduated  from  o°  to  360°  in  both 
directions,  that  is,  both  to  the  left  and  right.  The  inner  graduation 
is  the  one  necessary  for  running  continuous  vernier,  as  will  be 
described  later,  and  should  be  on  all  mining  transits.  The  gradua- 
tion on  the  outer  circle  is  of  no  particular  value  unless  it  should 
some  time  prove  desirable  to  work  with  the  telescope  reversed. 


FIG.    Q.      METHODS   OF   GRADUATING   HORIZONTAL   CIRCLES. 

The  smaller  or  inner  circle  of  graduations  is  perhaps  the  most 
useful  for  general  mining  practice  there  is.  As  will  be  noticed  the 
outer  circle  of  figures  is  graduated  for  the  continuous  vernier  as 
already  mentioned,  wrhile  the  inner  one  is  graduated  for  reading 
courses  direct,  that  is  the  graduations  run  both  ways  from  o  at 


iS 


COAL  MINE  SURVEYING 


both  the  north  and  south  ends  terminating  at  90  on  both  the  east 
and  west.  Where  it  is  the  practice  (as  for  instance  in  the  engineer- 
ing department  of  the  Consolidation  Coal  Co.)  for  the  instrument- 
man  to  read  both  the  azimuth  and  bearing  of  all  sights,  this  style  of 
graduation  is  essential.  It  is  also  more  convenient  when  turning 
rights  and  lefts  as  is  practised  in  railroad  surveys,  and  is  useful  as 
well  in  laying  out  right-angle  work  in  the  mine. 

By  referring  to  the  accompanying  illustration,  Fig.  10,  which  repre- 
sents the  method  of  lettering  a  compass  it  will  be  observed  that  the 
east  and  west  marks  are  the  reverse  of  what  they  actually  are.  Thus 

when  the  needle  is  pointing 
to  the  north,  as  shown  in  the 
left-hand  illustration,  east 
would  be  to  the  right  where 
the  W  is  and  west  to  the  left 
where  the  E  is.  This  is  done 
as  a  matter  of  convenience 
and  to  avoid  the  possibility 
of  error  in  reading. 

Referring  to  the  right-hand 
illustration  in  Fig.  10,  let  it 
be  assumed  that  the  line  of 
sight  being  taken  is  in  the 
direction  indicated,  the 

^^Q   Pointing    tO    the    tTUC 

north,  as  is  also  shown.  It  is 
clear  that  the  line  of  sight  is  to  the  left  or  northwest,  which  is  also 
the  reading  shown,  that  is,  the  needle  falls  between  the  N  and  the  W 
indicating  a  northwest  reading;  were  the  lettering  placed  in  the 
reverse  order  as  is  the  case  on  most  small  pocket  compasses,  the 
reading  would  have  been  northeast,  which  would  of  course  have  been 
wrong.  This  same  condition  applies  to  vernier  readings  taken  on 
the  horizontal  circle  of  a  transit  and  explains  why  the  graduations 
are  arranged  as  they  are. 


COM.\6C 

FIG.    10.  METHOD  OF  LETTERING  THE  COMPASS. 


THE  LEVEL 

As  has  already  been  mentioned,  the  level  is  an  instrument  for 
determining  relative  elevations.  When  properly  adjusted,  it  turns 
in  a  horizontal  plane  showing  exactly  the  same  elevation  in  whatever 
direction  the  telescope  may  be  pointed. 


SURVEYING  INSTRUMENTS  AND  ACCESSORIES  19 

The  accompanying  illustration,  Fig.  u,  shows  a  typical  Keuffel 
&  Esser  level,  one  of  the  popular  makes  in  the  market  to-day.  As 
will  be  observed,  the  different  parts  are  numbered,  the  subjoined 
table  giving  the  technical  name  for  each.  It  will  also  be  noted  that 
many  of  the  names  for  the  different  parts  correspond  to  those 
already  given  in  the  description  of  the  transit.  Most  of  the  con- 
cealed parts,  as,  for  instance,  the  interior  of  the  telescope,  and  the 
half-ball  socket  joint  are  the  same  as  in  the  transit,  and  reference  to 
the  line  drawing,  Fig.  6,  may  be  made  for  determining  these. 

The  Telescope. — At  i,  in  Fig.  n,  is  the  objective  of  the  telescope 
and  7  the  eyepiece  tube,  with  the  micrometer  focusing  screw  and 
adjusting  nut  at  20  and  19  respectively.  The  adjusting  screws  for 
correcting  any  error  in  the  cross  wires  are  shown  at  6.  One  of  the 
wye  yokes  is  shown  at  2,  3  and  4,  the  spring  lock  being  at  2,  the 
spring  contact  at  3,  and  the  yoke  catch  at  4.  These  yokes  may  be 
thrown  back,  and  the  telescope  lifted  out,  as  is  necessary  when  mak- 
ing adjustments.  At  5  is  shown  the  rack  and  pinion  thumb-screw 
for  focusing  the  telescope. 

The  level  bubble  tube  by  which  the  instrument  is  leveled  up  is 
shown  at  12.  It  is  first  set  across  one  opposite  pair  of  leveling  screws 
and  leveled,  and  then  swung  across  the  other  two,  and  the  operation 
repeated  until  the  bubble  is  exactly  level  in  whatever  position  the 
telescope  may  be  turned.  The  bubble  is  always  adjusted  with  re- 
gard to  the  line  of  sight  through  the  instrument,  such  adjustment 
being  effected  by  means  of  three  screws,  one  on  the  left  at  n  and  two 
on  the  right  at  24  and  25. 

Level  Bar  and  Telescope  Wye.— The  level  bar  is  shown  at  13, 
at  each  end  of  which  are  the  two  wyes  for  supporting  the  telescope. 
At  9  and  10,  and  22  and  23  are  shown  the  two  sets  of  adjusting  nuts 
for  making  corrections  in  the  elevation  of  the  wyes.  At  21  is 
the  stop  lever  to  prevent  the  telescope  from  revolving  around  in 
the  wyes. 

The  tangent  clamp  screw  is  shown  at  26,  and  the  collar  at  14. 
This  is  useful  occasionally  when  it  is  necessary  to  take  a  series  of 
sights  on  a  rod  at  some  particular  point,  or  in  some  certain  direction. 
The  tangent  screw  for  slowly  shifting  the  range  of  the  telescope  is 
shown  at  27. 

The  leveling  head  of  the  instrument  is  shown  at  15,  with  one  of 
the  leveling  screws  at  16,  and  a  leveling  screw  shoe  at  17.  This 
part  of  the  instrument  corresponds  almost  identically  with  that 
already  described  for  the  transit.  At  28  is  the  half  ball  socket  joint 


20  COAL  MINE  SURVEYING 

which  allows  the  horizontal  line  of  the  instrument  to  be  at  any  angle 
with  the  tripod  head,  shown  at  18.  The  top  of  the  tripod  is  seen 
below. 

LEVEL  ROD 

In  the  illustration,  Fig.   12,  is  shown  one  of  the  modern  mine 
level  rods,  a  product  of  the  Keuffel  &  Esser  Co.    This  rod  is  a  trifle 


FIG.    II.     TYPICAL  WYE  LEVEL  USED  ON  MINE  WORK. 


I.  Objective 
2.  Spring  lock  lever  for  wye 
yokes 
3.  Spring  contact 
4.  Wye  yoke  catch 
5.  Rack  and  pinion  thumb- 
screw  (focusing  screw) 
6.  Adjusting      screws      for 
cross-wires 
7.  Eyepiece  tube 
8.  Telescope  wye 
9.  Wye  adjusting  nut 
10.  Wye  adjusting  nut 
u.  Level   bubble  adjusting   nut 
12.  Level  bubble  tube 
13.  Level  bar 
14.  Tangent  clamp  collar 

15.  Leveling  head 
16.  Leveling  screw 
17.  Leveling  screw  shoe 
1  8.  Tripod  head 
19.  Adjusting  nut 
20.  Micrometer          focusing 
screw 
21.  Stop-lever 
22.  Wye  adjusting  nut 
23.  Wye  adjusting  nut 
24.  Level    bubble    adjusting 
nut 
25.  Level    bubble    adjusting 
nut 
26.  Tangent  clamp   screw 
27.  Tangent  screw 
28.  Half  ball  socket  joint 

over  3  ft.  long,  overall,  and  can  be  extended  to  5  ft.  This  is  one  of 
the  most  common,  in  fact  it  might  be  said,  practically  the  only  type 
of  level  rod  used  on  mine  work. 


W 


SURVEYING  INSTRUMENTS  AND  ACCESSORIES  21 

Referring  to  Fig.  12  which  shows  the  completed  rod,  the  target 
will  be  observed  near  the  top,  and  a  set  screw  arrangement  for  clamp- 
ing the  rod  in  position  when  extended  is  shown  at  A.  Near  the 
bottom  of  the  rod  at  D  is  a  metal  sleeve  which  is  a  part  of  the  tele- 
scoping arrangement,  being  attached  to  the  back  half  of  the  rod  and 
sliding  along  the  front  part.  The  extreme  bottom  of 
the  rod  at  F  is  bound  with  a  substantial  iron  ferrule 
which  provides  against  any  important  wear  and  con- 
sequent inaccuracy  to  this  much  abused  part.  As 
will  be  noted,  the  graduations  on  the  rod  extend  from 
zero  at  the  bottom  to  3  ft.  at  the  top,  the  decimal 
system  of  tenths  and  hundredths  of  course  being  used 
instead  of  inches.  The  tenths  of  feet  are  plain  black 
figures,  the  even  foot  marks  being  indicated  by  a 
slightly  larger  figure  in  red. 

In  Fig.  13  is  shown  an  enlarged  view  of  the  target. 
The  target  is  simply  a  circular  disk  with  an  oblong 
portion  removed  in  the  center.  The  zero  of  the  tar- 
get is  indicated  by  sharp  changes  in  the  color  arrange- 
ment in  order  to  bring  this  out  prominently,  the  two 
colors  most  commonly  used  being  red  and  white,  as 
indicated  on  the  half-tone  by  R  and  W  respectively. 
To  the  right  at  C  will  be  noted  a  slit  and  small  cir- 
cular hole  in  the  target  which  is  used  on  under- 
ground work  only.  The  rodman,  by  holding  his  light 
behind  the  target  and  allowing  it  to  shine  through 
this  slit  brings  out  a  sharply  defined  isolated  mark 
which  the  instrumentman  cannot  fail  to  readily 
locate. 

A  modern  helpful  addition  to  the  target  is  the  device 
shown  at  B.     This  is  a  slow  motion  arrangement  by 
which  the  target  is  more  conveniently  set  at  the  exact 
joint.     On  mine  work,  particularly,  the  damp  atmosphere  results  in 
a  certain  swelling  of  the  wood  so  that  the  target  is  often  stiff,  and 
with  the  rodman  endeavoring  to  hold  the  light  with  one  hand  and 
move  the  target  an  infinitesimal  amount,  while  at  the  same  time 
holding  the  rod  plumb,  the  results  are  often  trying  on  the  patience 
of  all  concerned 

To  obtain  the  reading  of  the  target,  set  at  any  point  on  the  rod, 
first  note  the  nearest  foot  mark  beneath  the  target  which  gives  the 
full  number  of  feet  in  the  required  reading.  Next  observe  what 


22  COAL  MINE  SURVEYING 

tenth  mark  on  the  rod  cuts  the  small  scale  on  the  target,  this  being 
i  in  the  illustration,  Fig.  13.  Finally,  note  the  reading  on  this  small 
scale  at  the  before  mentioned  tenth  mark  which  in  this  case  is  .07, 
or  to  be  more  exact  about  .073  if  it  is  desired  to  carry  the  readings 
to  thousandths.  The  even  foot  mark  in  this  illustration  obviously 
occurs  between  the  9  and  the  i  on  the  rod  and  is  therefore  covered 


FIG.    13.      DETAIL   OF  .TARGET   ON   THE   MINE    LEVEL   ROD. 

by  the  target.     Assuming  this  to  be  2  we  then  have  a  reading  of 
2.173  ft. 

When  it  becomes  necessary  to  use  "long  rod"  the  target  is  run 
up  as  far  as  it  will  go,  which  is  to  the  3-ft.  mark  (see  Fig.  12)  and 
clamped  tight.  The  set  screw  A  is  then  loosened,  the  rod  extended 
to  the  required  height  and  the  set  screw  again  clamped.  On  the 
back  of  the  rod  will  be  found  another  small  scale  the  same  as  in 
the  target  and  the  reading  is  taken  in  the  same  way,  except  that  it 
is  made  from  the  top  down,  instead  of  from  the  bottom  up  as  in 
the  first  case. 


SURVEYING  INSTRUMENTS  AND  ACCESSORIES  23 

STEEL  TAPES 

Unusual  conditions  obtain  in  mine  surveying  that  accentuate  the 
possibility  of  error  in  chaining  and  the  difficulty  in  obtaining  reliable 
results.  Where  the  average  engineer  doing  outside  work  is  liable  to 
lose  his  temper  should  a  thoughtless  chainman  inadvertently  drag 
the  tape  through  mud  or  water,  the  inside  man  has  long  since  ac- 
cepted this  as  a  matter  of  course.  The  outside  engineer  is  doing 
his  work  in  broad  daylight  where  the  most  casual  attention  only  is 
required,  while  the  inside  man  is  working  in  a  pitch  blackness  where 
kinks  in  the  tapes  are  seldom  found  until  too  late,  where  it  is  difficult 
to  tell  if  the  tape  swings  free  and  truly  on  the  line,  and  where  the 
limited  working  room  requires  constant  winding  and  unwinding  of 
the  tape. 

As  a  result  of  these  abnormal  conditions,  a  distinctive  style  of  tape 
has  been  developed  on  mining  work.  This  is  the  ribbon  or  flat  wire 
tape,  as  shown  in  Fig.  14.  These  tapes  are  ordinarily  |  in.  wide, 


FIG.    14.      FLAT    WIRE    OR    RIBBON    TAPE    USED    IN    MINE    SURVEYING. 

and  usually  about  300  ft.  long.  Some  engineers  prefer  even  a 
lighter  tape  only  -jV  in.  wide,  and  have  them  in  lengths  up  to  500 
ft.,  and  even  more  on  occasions. 

As  noted  in  Fig.  14,  the  graduations  on  this  type  of  tape  are 
stamped  on  brass  sleeves,  which  have  been  either  clamped  or 
soldered  onto  the  tape.  The  figures  indicating  the  number  of  feet 
from  zero  are  noted,  together  with  a  line  and  a  notch  on  each  side 
at  the  exact  point.  These  graduations  are  usually  put  on  every  5 
ft.  although  sometimes  3  ft.  is  used.  The  less  there  are,  the  more 
accurate  the  tape  will  be,  and  the  less  expensive.  The  $-f  t.  gradua- 
tions are  therefore  probably  the  best,  particularly  as  the  experienced 
man  soon  becomes  so  accustomed  to  this  style  that  he  can  handle  it 
as  rapidly  as  the  3-ft.  graduations. 

These  tapes  are  made  of  the  toughest  flexible  steel  ribbon,  carefully 
tempered,  so  as  to  withstand  breaking  under  the  hardest  usage. 
They  are  graduated  according  to  the  standard  of  the  National 
Bureau  of  Standards,  and  are  correct  at  62°  F.  On  the  ordinary 
mine  surveying  work,  it  is  not  necessary  to  make  corrections  for 
temperature  or  the  tension  on  the  tape.  Another  advantage  of  these 


24  COAL  MINE  SURVEYING 

tapes  is  the  facility  with  which  they  can  be  repaired.  This  latter 
contingency  is  one  that  must  be  anticipated  and  prepared  for  even 
with  the  well- trained  corps.  The  following  method  of  effecting  re- 
pairs, selected  from  a  number  appearing  in  Engineering  News 
several  years  ago,  will  be  found  convenient: 

Small  pieces  of  copper,  or  tin  from  an  old  tin  can,  are  cut  into  strips 
say  f  in.  wide,  and  then  cut  crosswise  of  a  size  that  will  lap  around 
the  tape  to  be  mended  and  just  about  meet  on  the  flat  side  of  the 
tape — not  the  edge.  These,  with  an  ordinary  candle,  a  piece  of  solder 
and  some  stick  flux  (both  of  which  latter  may  be  obtained  from  any 
electric  plant),  a  small  flat  file  and  a  pair  of  small  nippers  complete 
the  outfit.  These  may  all  be  put  into  a  small  sack  and  carried 
in  the  pocket,  where  they  are  always  ready,  and  when  a  tape  is  broken 
it  can  be  repaired  in  the  field  in  five  minutes,  thus  saving  time  and 
the  inconvenience  of  using  a  broken  tape,  or  going  a  long  distance  to 
get  it  repaired. 

One  end  of  the  tape  is  used  as  a  "form"  about  which  the  clasp 
is  bent.  The  tape  and  clasp  should  be  brightened  with  the  file 
before  the  clasp  is  bent  into  the  proper  shape.  When  the  tape  is 
put  together,  clamp  the  clasp  tightly  about  it  with  the  nippers; 
heat  in  the  candle  and  put  sufficient  flux  on  to  run  under  the  clasp; 
then  hold  the  solder  on  the  splice  until  it  has  melted  and  run  under 
and  about  the  clasp;  allow  it  to  cool  without  movement  of  the  ends 
of  the  tape  and  the  work  is  done. 

The  above  " outfit"  does  not  weigh  over  a  pound,  and  is  all  that 
will  be  needed,  thus  doing  away  with  a  machine  shop  for  doing  a  little 
tape  splicing.  Tapes  can  easily  be  mended  in  this  manner  so  that 
they  will  hold  all  a  man  can  pull.  The  stick  flux  is  much  more 
convenient  than  a  bottle  of  acid  and  zinc,  as  it  cannot  be  broken, 
and  does  the  work  as  well  as  the  acid. 


PLUMB  BOBS 

Plumb  bobs  are  used  in  mine  work  for  giving  sights,  but  more 
particularly  for  setting  up  the  instrument.  There  is  little  choice  in 
the  matter  of  bobs,  Figs.  15  and  16  showing  the  two  principal  varie- 
ties on  the  market.  Fig.  15  shows  the  standard  type  of  plumb  bob 
most  commonly  used.  It  conforms  to  the  general  plan  adopted  since 
plumb  bobs  were  first  made,  and  has  successfully  maintained  its 
leadership  to  the  present  day.  Fig.  16  shows  a  relative  new  type  of 
bob  made  of  nickel  steel  instead  of  brass,  as  is  the  one  shown  in  Fig. 


SURVEYING  INSTRUMENTS  AND  ACCESSORIES  25 

15.  This  new  type  is  much  preferred  by  many  colliery  engineers  who 
claim  that  it  offers  much  less  resistance  to  the  rapid  air  current  pre- 
vailing in  some  parts  of  our  mines,  and  it  is  much  easier  to  work  with 


PIG.  15  FIG.  16 

under  such  conditions.  Some  also  find  the  cylindrical  form  more 
convenient  for  handling  as  compared  with  the  conical  form  of  the 
old  style  bob. 


CHAPTER  III 
CARE  OF  INSTRUMENTS 

Note. — Abstracted  from  publications  of  the  C.  L.  Berger  Co. 

The  first  requisites  in  the  outfit  of  the  engineer  are  the  engineering 
instruments,  transit,  level,  rods,  tapes,  etc.  The  perfect  instru- 
ment becomes  a  part  of  the  operator,  like  his  skillful  hand  and 
educated  brain,  and  in  like  manner  can  be  depended  upon.  To 
make  such  an  instrument  requires  long  experience,  great  skill,  and 
the  use  of  delicate  and  expensive  machinery.  Some  of  the  machinery, 
like  the  automatic  dividing  engines,  require  many  years  of  construc- 
tion and  perfection. 

Before  purchasing  of  any  particular  manufacturer,  consult 
engineers  of  long  experience  and  high  standing,  as  to  the  merits  of 
the  instruments  of  different  makers.  A  certain  high  quality  in 
instruments  is  requisite  to  do  good  work  quickly.  Do  not  let  the  mat- 
ter of  a  few  dollars  in  the  difference  in  price  influence  you  to  purchase 
that  for  which  you  may  soon  have  to  apologize.  Patronize  a  firm 
that  has  a  well-established  record  for  fair  dealing  and  for  reliable 
work  that  will  give  you  full  value  for  your  money. 

The  perfect  instrument  is  so  made  that  every  part  is  designed  with 
reference  to  every  other  part.  Strength,  weight,  rigidity,  and  sta- 
bility under  wind  pressure,  are  carefully  considered,  as  well  as  in  the 
form,  the  material,  the  life  and  the  action  of  the  tripod,  movement, 
centers,  bearings,  leveling  and  tangent  screws,  telescope  slides,  the 
power  and  clearness  of  the  telescope,  the  sensitiveness  of  the  levels, 
the  accuracy  of  the  graduations,  and  the  simplicity  of  the  manipula- 
tion. The  instrument  must  be  constructed  so  that  it  can  be  used 
in  all  climates  and  under  all  conditions,  and  when  exposed  to  wind 
pressure  all  tremor  and  vibration  is  eliminated,  and  lines  and  angles 
can  be  laid  out  and  measured  correctly  without  any  anxiety  as 
regards  results  and  nervousness  therein  on  the  part  of  the  manipu- 
lator. 

IN  GENERAL 

Do  not  allow  the  legs  of  your  tripod  to  play  loose  on  the  tripod 
head;  keep  nuts  and  bolts  always  well  tightened  up  against  the 

26 


CARE  OF  INSTRUMENTS  27 

wood.  Examine  the  shoes  from  time  to  time,  and  sharpen  them  if 
necessary,  also  screw  the  shoes  tight,  if  wear  and  tear  loosen  them. 
Be  sure  your  instrument  is  well  secured  to  its  tripod  before  using  it. 
Bring  all  four  leveling  screws  to  a  seat  before  shouldering  instru- 
ment. Let  the  needle  down  upon  its  pivot  as  gently  as  possible, 
and  allow  it  to  play  only  when  in  use;  if  too  far  out  from  its  course, 
check  movements  of  needle  carefully  by  means  of  lifter.  Never 
permit  playing  with  the  needle,  especially  not  with  knives,  keys, 
etc.  Be  sure  to  arrest  the  needle  after  use,  and  screw  it  well  up 
against  the  glass  cover  before  shouldering  the  instrument. 

Do  not  clean  the  glass  cover  or  the  lenses  with  a  silk  handkerchief; 
breathe  over  the  compass  glass  and  reading  lens  if  one  is  used,  after 
cleaning.  To  clean  the  object  glass  and  the  lenses  use  a  fine  camel- 
hair  brush.  If  dust  or  sticky  or  fatty  matter  cannot  be  removed  with 
the  brush,  take  an  old  clean  piece  of  soft  linen,  and  carefully  wipe  it 
off.  Do  not  unscrew  the  object  glass  unnecessarily  as  this  is  apt  to 
disturb  the  adjustment  of  line  of  collimation.  The  lens  nearest 
the  eye  of  eyepiece,  as  well  as  the  front  side  of  the  object  glass, 
need  careful  brushing  with  fine  brush  from  time  to  time. 

If  dust  settles  on  cross  hairs  and  becomes  troublesome,  unscrew 
the  eyepiece  and  object  glass,  and  gently  blow  through  the  telescope 
tube,  cover  up  both  ends  and  wait  a  few  minutes  before  inserting  the 
eyepiece  and  object  glass.  Be  sure  to  have  the  object  glass  cell 
screwed  well  up  against  its  shoulder,  and  then  examine  the  adjust- 
ment for  the  line  of  collimation.  Do  not  grease  the  object  slide  of 
telescope,  or  screws  that  are  exposed  to  dust;  use  a  stiff  toothbrush 
to  clean  slides  or  threads  if  dusty. 

If  the  focussing  slide  seems  to  work  too  hard,  everything  else  being 
right,  it  is  generally  caused  by  the  lubricant  on  the  pinion  hardening 
in  cold  weather,  and  the  same  cause  may  also  make  the  focussing 
slide  work  too  freely  in  hot  weather  by  softening,  i.e.,  when  not  stay- 
ing in  place  when  in  a  vertical  position.  Fretting  of  the  focussing 
slide  is  usually  due  to  the  inrush  of  air  carrying  dust  and  grit  when 
slide  is  being  run  out  causing  momentarily  a  rarefied  space.  This 
may  be  prevented  by  wrapping  a  piece  of  chamois  skin  over  the  barrel 
in  shape  of  tubular  form  and  fasten  by  means  of  rubber  bands  or 
sewing.  In  an  emergency  fine  watch-oil  may  be  used  to  grease 
the  slide  should  it  continue  to  fret,  until  the  instrument  can  be 
sent  to  the  maker.  In  case  of  rain  during  non-use,  place  the  telescope 
vertical,  object  end  up,  and  no  water  can  enter  the  telescope. 

Never  use  emery  in  any  form  about  any  part  of  a  transit  or  a  level, 


28  COAL  MINE  SURVEYING 

whether  tangent  screws,  slides  or  centers.  If  anything  must  be 
used,  a  very  little  powdered  pumice-stone  mixed  with  fine  watch- 
oil  is  all  that  is  advisable,  and  after  grinding,  then  clean  thoroughly. 
The  uninitiated  are  advised  to  do  no  grinding  whatever.  As  a 
rule  more  harm  than  good  comes  to  the  instrument.  It  is  only  in 
case  of  emergency  that  such  heroic  treatment  should  be  resorted  to. 
When  cleaning  the  slide  and  inside  of  main  tube  great  care  must  be 
taken  not  to  break  the  wires. 

To  clean  the  threads  of  leveling  or  tangent  screws  when  working 
hard,  use  a  stiff  toothbrush  to  remove  all  dust,  then  apply  a  little  oil, 
and  work  the  screw  in  and  out  with  alternate  brushing  to  remove  dirt 
and  all  oil  until  it  moves  perfectly  free  and  smooth. 

Screws  for  the  adjustment  of  cross  hairs  should  not  be  strained  any 
more  than  necessary  to  insure  a  firm  seat;  all  straining  of  such  screws 
beyond  this  simply  impairs  the  accuracy  of  instrument  and  reliability 
of  adjustment. 

When  in  the  field  always  carry  a  Gossamer  water-proof  to  put  over 
the  instrument  in  case  of  a  shower  or  dust  cloud.  On  reaching  office, 
after  use  of  instrument,  dust  it  off  generally  with  another  fine  brush; 
examine  the  centers  and  all  other  principal  movements  to  see  if 
they  run  perfectly  free  and  easy,  and  oil  them  if  necessary;  also 
examine  the  adjustments.  This  will  save  expense  and  many  hours  of 
vexation  in  the  field. 

In  field  use,  an  instrument  has  to  be  necessarily  exposed  to  the  heat 
of  the  sun,  and  to  the  action  of  dust  and  water;  all  of  these,  however, 
singly  or  combined,  have  a  tendency  to  affect  its  accuracy  and 
endurance.  While  good  instruments  are  designed  to  guard  against 
injuries  resulting  from  exposure  of  this  kind,  yet  glaring  abuses, 
such  as  to  allow  it  to  stand  for  hours  in  the  hot  sun,  etc.,  without  a 
covering  or  shelter  of  some  sort,  may  often  lead  to  a  permanent  injury 
to  its  most  vital  parts.  To  preserve  the  finer  qualities  of  an  in- 
strument, viz.,  the  telescope  slide,  the  lenses,  the  edge  of  the  gradua- 
tion and  verniers,  the  centers,  etc.,  any  undue  unequal  expansion  of 
the  different  parts  should  be  prevented.  A  bag  thrown  over  the  in- 
strument when  not  in  use,  or  any  shelter  that  can  be  had,  is  to  be 
recommended.  While  in  use,  an  umbrella  or  screen  held  over  it  will 
insure  greater  permanency  of  its  adjustments,  and  the  results  ob- 
tained will  be  more  accurate  and  uniform  than  when  carelessly 
exposed. 

To  protect  an  instrument  from  the  effects  of  salt  water,  a  fine  film 
of  watch-oil  rubbed  over  the  exposed  parts  will  often  prevent  the 


CARE  OF  INSTRUMENTS  29 

appearance  of  oxyd.  To  remove  such  oxyd-spots  as  well  as  possible, 
apply  some  watch-oil  and  allow  it  to  remain  for  a  few  hours,  then  rub 
dry  with  a  soft  piece  of  linen.  To  preserve  the  outer  appearance  of 
an  instrument,  never  use  anything  for  dusting  except  a  fine  camel's 
hair  brush.  To  remove  water  and  dust  spots,  first  use  the  camel's 
hair  brush,  and  then  rub  off  with  fine  watch-oil,  and  wipe  dry;  to 
let  the  oil  remain  would  tend  to  accumulate  dust  on  the  instrument. 

CARE  OF  THE  CENTERS  AND  GRADUATIONS    ,. 

If  it  is  found  that  the  centers  do  not  revolve  freely,  as  is  often  the 
case  after  exposure  to  extremes  of  temperature,  take  the  instrument 
apart  and  proceed  as  follows: 

Take  a  fine  camel's  hair  brush,  and  with  it  clean  the  graduation, 
the  verniers  and  the  inner  part  of  the  instrument,  but  do  not  rub 
the  graduation,  especially  not  its  edge.  Then  take  a  stick  of  about 
the  same  taper  as  the  inner  center,  wrap  some  wash-leather  slightly 
soaked  in  fine  oil  around  it,  and  clean  the  insides  of  the  sockets  as 
carefully  as  possible  and  then  wrap  a  fresh  piece  without  oil  around 
the  stick  and  clean  dry.  Proceed  similarly  with  the  centers  and 
their  flanges. 

Before  applying  fresh  and  pure  watch-oil,  care  should  be  taken 
that  not  a  particle  of  dust  or  other  foreign  matter  is  left,  in  the  sock- 
ets, on  the  centers,  or  on  the  graduation.  This  caution  having  been 
taken,  the  fresh  oil  should  be  well  distributed  on  all  the  bearing  parts. 
It  will  be  well  to  also  examine  the  arm  of  the  clamp  screw  of  the  circle 
and  telescope  axis,  and  if  necessary  clean  by  removing  washer. 
After  the  instrument  is  thoroughly  cleaned  and  oiled,  the  nuts  and 
springs  screwed  back  to  a  firm  seat,  the  instrument  must  turn 
perfectly  free  and  yield  at  the  slightest  touch  of  the  hand. 

To  remove  dirt  and  oxyd  that  may  have  accumulated  on  the 
surface  of  a  solid  silver  graduation,  apply  some  fine  watch-oil,  and 
allow  it  to  remain  for  a  few  hours;  take  a  soft  piece  of  old  linen  and 
slightly  rub  until  dry,  but  without  touching  the  edge  of  the  gradua- 
tions. If,  after  cleaning,  the  solid  silver  surface  should  show  alter- 
nately brighter  spots,  which  would  interfere  somewhat  with  the 
accurate  reading  of  the  graduation,  barely  moisten  the  finger  with 
vaseline  and  apply  the  same  to  the  surface;  then  wipe  the  finger  dry 
and  lightly  rub  it  once  or  twice  around  the  graduation.  Avoid 
touching  the  edges  as  much  as  possible.  Such  cleaning,  however, 
must  only  be  resorted  to  when  absolutely  necessary,  and  then  only 


3o  COAL  MINE  SURVEYING 

with  the  greatest  care,  as  it  is  too  apt  to  reduce  the  minuteness  of 
the  graduation,  and  spoil  its  fine  appearance.  If,  after  such  cleaning, 
dirt  and  grease  has  accumulated  on  the  inner  edge  of  the  graduation 
and  verniers,  gently  wipe  clean  before  restoring  the  vernier-plate 
to  its  place.  Remember,  also,  that  the  centering  of  the  graduations 
of  the  circle  and  verniers  is  a  most  delicate  adjustment  to  make. 
These  should  never  be  unscrewed  from  their  flanges  by  anybody 
except  a  maker. 

TELESCOPE  LENSES 

As  dust  and  moisture,  as  well  as  perspiration  from  the  hands,  will 
settle  on  the  surface  of  the  lenses  of  a  telescope,  it  becomes  necessary 
that  they  should  be  cleaned  at  times.  A  neglect  to  keep  the  lenses 
free  from  any  film,  scratches,  etc.,  greatly  impairs  the  clear  sight 
through  the  telescope.  To  remove  the  dimness  produced  by  such  a 
film,  proceed  thus: 

Brush  each  lens  carefully  with  a  camel's  hair  brush,  wipe  gently 
with  a  clean  piece  of  chamois  leather  moistened  with  alcohol,  and 
wipe  dry  using  a  clean  part  of  the  chamois  skin  on  every  portion  of 
the  lens,  to  avoid  grinding  and  scratching.  When  perfectly  trans- 
parent brush  again  to  remove  any  fiber  that  may  adhere  to  the  lens. 
The  tubes  in  which  the  lenses  fit  should  be  brushed,  and  if  damp 
should  should  be  dried;  this  done,  restore  each  lens  to  its  original 
place  as  marked.  To  remove  dampness  in  the  main  tube  of  the 
telescope,  take  out  the  eyepiece,  cover  the  open  end  with  cloth  and 
leave  the  instrument  in  a  dry  room  for  some  time. 

If  an  instrument  has  been  exposed  to  a  damp  atmosphere,  or  water 
has  penetrated  the  telescope,  moisture  may  settle  between  the  crown 
and  flint  glass  of  which  the  object  glass  is  composed.  If  such  is  the 
case  expose  the  instrument  to  the  sun  for  a  few  hours,  but  if  in  the 
winter,  leave  it  in  a  warm  room  some  distance  from  the  stove,  the 
moisture  will  then  generally  evaporate.  However,  if  not  successful, 
unscrew  the  object  glass  from  the  telescope,  and  heat  it  slightly  over 
a  stove  or  open  fire.  If  a  film  settles  between  these  glasses  nothing 
can  be  done  except  sending  the  instrument  to  the  maker.  The 
two  glasses  form  one  lens  only  and  must  not  be  disturbed,  as  upon 
their  relation  to  each  other  the  definition  and  achromaticity  of 
the  telescope  depends.  Much  depends  also  on  the  stability,  with 
which  these  lenses  are  mounted  in  their  cell,  as  any  looseness  between 
them  or  the  cell  will  affect  the  adjustment  of  line  of  collimation. 


CARE  OF  INSTRUMENTS  31 

Of  course,  if  at  any  time  the  object  glass  has  been  unscrewed  from 
the  telescope,  this  latter  adjustment  must  again  be  verified  before 
the  instrument  is  used. 

When  the  object  glass,  or  telescope  is  returned  after  the  cleaning  or 
cementing  of  its  lenses,  the  cross-wire,  spirit  level,  and  vertical  arc 
adjustments  of  the  instrument  will  require  a  thorough  verification 
before  it  should  be  used.  In  case  the  whole  instrument  has  been 
sent  to  the  maker,  these  adjustments  are  attended  to  by  him.  If  the 
object  glass  has  been  cemented,  the  telescope  should  be  watched  for 
a  year  to  see  that  there  is  no  distortion  of  the  image.  If  there  is  a 
distortion,  it  will  indicate  that  the  object  glass  has  been  too  tightly 
fitted,  of  which  fact  the  makers  sKould  be  informed,  as  also  whether 
after  cementing  the  object  glass  the  instrument  retains  its  cross- 
wire  adjustment  the  same  as  before.  If  the  cross-wire  adjust- 
ments have  to  be  more  frequently  made  than  before  the  lenses  were 
cemented,  it  indicates  that  the  object  glass  is  not  tightly  fitted  to  its 
cell;  and  if  such  is  the  case  it  should  be  returned  and  more  tightly 
fitted,  after  a  lapse  of  about  ten  or  twelve  months,  when  the  cement 
will  have  sufficiently  hardened  to  allow  of  a  tighter  fit. 

LUBRICATING 

An  instrument  used  in  a  tropical  or  semi-tropical  country,  or 
during  the  warm  season  in  a  northern  latitude,  requires  more  fre- 
quent cleaning  and  oiling  than  in  the  more  temperate  climes  and 
seasons;  but  so  long  as  an  instrument  works  well  and  the  centers 
revolve  freely,  it  is  best  not  to  disturb  it. 

The  centers  of  a  transit  should  always  be  lubricated  with  fine  watch- 
oil  only,  and  after  a  careful  cleaning;  never  apply  fresh  oil  before 
thoroughly  wiping  off  old  grit  and  oil.  Rendered  marrow  is  a  most 
excellent  lubricant  for  instruments  made  of  brass  and  the  many  kin- 
dred alloys  of  copper  and  tin.  In  the  varying  climes  of  our  north- 
ern latitudes  this  lubricant  becomes  rigid  in  cold  weather,  and 
an  instrument  so  treated  will  often  become  unmanageable  in  the 
field.  Its  application,  particularly  to  the  centers  of  a  transit,  is 
therefore  restricted  to  the  warmer  zones.  The  use  of  watch-oil  for 
the  finer  parts  of  an  instrument,  involving  freedom  of  motion,  is 
imperative  in  our  latitudes. 

Many  parts  of  an  instrument,  especially  those  whose  metal  com- 
positions are  closely  related  to  each  other,  may  sometimes  cause 
trouble  if  simply  oiled.  If  they  begin  to  fret  and  grind,  but  are  other- 


32  COAL  MINE  SURVEYING 

wise  free  from  grit,  etc.,  the  judicious  application  of  a  little  marrow 
may  prove  beneficial,  but  it  should  be  cleaned  off  again  as  much  as 
possible.  The  rack  and  pinion  motion  and  the  telescope  clamp 
should  always  be  greased  with  marrow,  but  the  clamp,  tangent  and 
leveling  screws,  should  receive  as  little  of  it  as  possible  in  the 
Northern  States. 

Vaseline,  not  having  as  great  a  tendency  to  solidify  under  similar 
circumstances,  may  prove  an  excellent  substitute  for  marrow,  and 
may  often  be  applied  to  level-centers,  where  watch-oil  would  not 
give  the  necessary  rigidity  in  the  use  of  the  more  ordinary  instru- 
ments, but  it  must  be  renewed  quite  often.  In  the  finer  class  of 
leveling  instruments,  the  centers  should  be  lubricated  with  oil  only, 
as  in  transits. 

A  great  deal  of  annoyance  is  caused  if  the  eyepiece  or  the  object 
slide  of  the  telescope  move  too  freely  in  their  tubes,  requiring  a  re- 
focussing  of  the  cross  wires  and  object  at  every  revolution  of  the 
telescope  in  altitude.  If  the  eyepiece  can  be  retained  in  its  socket, 
with  sufficient  friction  to  keep  it  f ocussed  to  the  cross  wires,  no  matter 
how  much  it  may  wabble  otherwise,  this  imperfection  (in  old  instru- 
ments) will  not  lead  to  any  inaccuracy,  but  if  there  is  not  sufficient 
friction  to  keep  it  focussed  to  the  wires,  a  little  rendered  tallow  or 
marrow  applied  to  its  bearing  surfaces  in  most  cases  will  remedy  this 
evil. 

Wabbling  in  the  object  slide,  however,  leading  to  inaccuracy  of 
collimation,  or  back-lash  in  its  rack  or  pinion  motion,  can  be  remedied 
only  by  a  maker;  but  if  the  object  slide  moves  too  freely  in  and  out 
of  its  tube  only,  this  may  be  remedied  by  applying  a  little  tallow 
to  the  bearing  parts  of  the  rack  and  pinion,  or  by  tightening  the 
screw  in  the  pinion-head.  If  not  entirely  successful,  a  thin  disk  made 
of  parchment,  or  a  thin  leather  washer,  both  greased  with  tallow, 
and  inserted  between  the  flanges  of  the  pinion-head  and  its  socket, 
will  insure  the  desired  result. 

These  latter  remarks  apply  to  transit  and  level  telescopes  of  the 
customary  design.  In  telescopes,  where  the  object  glass  is  mounted 
permanently  to  the  telescope  tube,  the  eyepiece  tube,  containing 
the  cross  wires,  becomes  the  slide  with  which  to  focus  the  object. 
Its  motion  must  be  in  a  line  parallel  to  the  optical  axis.  Any 
wabbling  in  this  eyepiece  slide  would  lead  to  inaccuracy  in  sighting 
through  the  telescope,  hence  it  requires  the  most  careful  treatment 
on  the  part  of  the  engineer. 


CARE  OF  INSTRUMENTS  33 

LEVEL  BUBBLES 

Spirit  levels  are  very  susceptible  to  the  least  change  in  tempera- 
ture, as  will  be  readily  seen  by  the  difference  in  the  length  of  its 
bubble  in  varying  temperatures.  Hence,  to  guard  against  inac- 
curacies from  this  source,  it  is  necessary  that  the  bubble  should 
lengthen  symmetrically  from  the  center  of  its  graduated  scale  (sup- 
posed to  be  put  on  by  the  maker),  and  that  both  of  its  ends  should 
be  read.  Sufficient  time  must  also  be  allowed  for  the  bubble  to 
settle  before  reading  is  made. 

The  fluid  ordinarily  used  for  levels  is  pure  alcohol,  and  requires, 
according  to  curvature,  diameter  and  length  of  tube  and  length  of 
bubble,  from  twenty  seconds  to  one  minute  to  attain  its  equilibrium. 
The  composition  fluid  used  in  some  levels  for  field  instruments 
requires  only  from  five  to  fifteen  seconds  of  time;  those  filled  with 
pure  ether,  a  few  seconds  only. 

A  great  source  of  error  in  spirit  levels,  however,  increasing  with 
their  greater  sensitiveness,  is  occasioned  by  an  unequal  heating  of 
the  level-tube,  as  the  bubble  will  always  move  toward  the  warmer 
spot  or  end,  thereby  imparting  to  the  instrument  an  inaccurate 
position.  This  must  be  attributed  to  a  changed  condition  in  the 
adhesiveness  of  the  fluid  in  the  level-tube,  and  not  to  a  change  in  the 
form  of  the  tube  itself.  Therefore,  to  guard  against  inaccuracy  re- 
sulting from  sudden  changes  of  temperature,  a  spirit  level,  while  in 
use,  should  be  protected  from  the  sun,  and  no  part  of  it  or  its  mount- 
ing should  ever  be  touched  with  bare  fingers;  neither  should  it  be 
breathed  upon,  nor  the  face  of  the  observer  come  too  close  to  it. 
For  this  reason,  in  the  finer  instruments  the  mountings  of  some  spirit- 
levels  are  cloth-finished,  and  if  the  levels  are  detachable  they  are 
provided  with  wooden  handles,  as  the  case  may  require,  and  glass 
covers  are  placed  over  them  whenever  deemed  necessary. 

If  at  any  time  during  the  progress  of  field-work  a  spirit  level 
has  been  improperly  exposed,  it  is  best  to  cover  it  with  a  cloth 
for  from  five  to  fifteen  minutes,  before  proceeding  with  further 
work. 

Mounting  Spirit  Levels. — To  prevent  any  undue  strain  and  change 
of  curvature  in  spirit  levels  used  in  astronomical  instruments,  they 
are  mounted  by  some  makers  in  wyes  and  are  protected  from  injury, 
or  inaccuracy  caused  by  the  breath  of  the  observer  and  other  air 
currents,  by  a  cover  of  glass  placed  over  them.  Such  a  mounting, 
while  most  suitable  for  such  delicate  levels,  would,  however,  require 
3 


34 


COAL  MINE  SURVEYING 


constant  attention  and  expose  a  spirit  level  to  breakage  in  field  in- 
struments. To  guard  against  this  danger  and  to  lessen  the  expense 
and  weight,  the  spirit  levels  for  field  instruments  are  mounted  in  a 
brass  tube;  but  owing  to  the  difference  existing  in  the  expansion  and 
contraction  of  glass  and  brass  at  different  temperatures,  a  spirit-level 
so  mounted  may  sometimes  become  loose,  involving  inaccuracy  and 
unreliability  of  adjustment. 

Upon  finding  that  the  adjustment  of  a  spirit  level  in  an  even  tem- 
perature is  not  as  stable  as  desirable,  the  level  fastenings,  tube, 
screws,  etc.,  should  be  examined,  to  see  if  any  of  them  are  loose.  If 
the  trouble  is  in  the  screws,  tighten  them  up;  but  if  the  spirit  level 
can  be  shifted  in  its  tube  by  a  touch  of  the  finger,  take  it  apart; 
soften  the  plaster  of  Paris  in  water,  and  remove  it  with  a  sharp 
pointed  stick  of  wood.  Cautiously  move  the  spirit  level  with  your 
finger,  at  first  only  a  trifle  to  and  fro,  increasing  the  length  of  stroke 
little  by  little,  until  it  can  be  safely  taken  out  without  breaking. 
Clean  thoroughly  and  then  cut  pieces  of  white  paper,  of  the  width  of 
the  radius  of  the  tube,  and  somewhat  shorter  than  the  length  of  the 
spirit  level,  but  longer  than  the  opening  in  the  brass  tube,  and  insert 
these  of  sufficient  quantity  at  the  bottom  of  the  brass  tube,  to  fill 
up  the  space  intervening  between  the  glass  and  the  brass  tube. 
The  uppermost  layer  of  paper  should,  however,  be  so  wide,  as  to 
envelop  the  spirit  level  up  to  the  opening  in  the  brass  tube. 

Now  insert  the  spirit  level,  taking  care  not  to  touch  the  glass  ends 
that  are  sealed  up,  arid  place  the  division  or  other  marks,  indicating 
where  the  level  has  been  ground  to  a  true  curvature,  uppermost  in 
the  brass  tube.  The  level  must  be  pushed  in  with  sufficient  friction 
to  prevent  slipping  in  the  tube,  yet  not  so  tight  as  to  cause  a  crack  at 
a  subsequent  low  temperature,  as  brass  will  contract  more  than 
glass.  No  part  of  the  spirit  level  should  touch  any  part  of  the  metal 
tube.  Now  prepare  some  plaster  of  Paris  with  water,  of  the  consis- 
tency of  paste,  and  pour  in  at  each  end  enough  to  fill  up  the  space 
between  the  end-pieces  and  the  glass,  stirring  it  sufficiently  to  make 
a  perfect  contact  by  it  and  the  glass  and  the  brass,  but  leaving  the 
spirit  level  ends  exposed.  Now  put  the  level  together,  and  adjust 
as  described  elsewhere. 

There  are  other  causes,  such  as  centers  and  flanges  that  have  been 
bent  by  falls,  etc.,  or  that  have  been  worn  out — unequal  expansion 
or  contraction  in  different  temperatures  of  the  metals  employed  in 
the  construction  of  an  instrument,  or  a  non-symmetrical  lengthening 
or  shortening  of  the  air-bubble  at  different  temperatures — all  of 


CARE  OF  INSTRUMENTS  35 

which,  singly  or  combined,  tend  to  impair  the  adjustment  of  spirit- 
levels  on  instruments. 

Being  assured  that  the  level  is  mounted  as  explained  above,  it 
is  advisable  not  to  meddle  too  frequently  with  the  adjustment. 
Though  it  may  appear  to  be  out  one  day,  it  may  be  in  perfect  adjust- 
ment other  days.  It  is  the  function  of  a  spirit  level  to  indicate  the 
changes  taking  place  in  an  instrument,  so  that  the  engineer  may  make 
proper  allowance  and  apply  his  corrections,  as  the  character  of  his 
work  may  require.  The  finer  an  instrument,  the  more  sensitive  the 
spirit  levels  must  be,  in  order  to  admit  of  corrections  to  arrive  at 
closer  results.  As  a  rule,  a  spirit  level  that  does  not  indicate  changes 
taking  place  in  an  instrument,  is  too  insensitive  for  the  character  of 
the  instrument,  and  in  many  cases  entirely  unfit  for  reasonably 
good  work. 

REPLACING  CROSS  WIRES 

Remove  the  reticule  frame  and  clean  it  of  all  foreign  matter; 
put  it  on  a  sheet  of  white  paper  with  tl^e  cuts  on  its  surface  upper- 
most. Prepare  a  little  shellac  by  dissolving  it  in  the  best  alcohol 
and  waiting  until  it  is  of  the  consistency  of  oil.  From  the  spider's 
cocoon  (those  from  a  small  black  wood-spider  preferred),  which  the 
engineer  has  prudently  secured  at  some  previous  time,  select  two  or 
three  webs,  each  about  2  in.  long  and  of  the  same  appearance. 
Attach  each  end  of  these  webs  to  a  bit  of  paper  or  wood  to  act  as 
weights,  and  immerse  them  in  water  for  five  or  ten  minutes.  Remove 
one  web  from  the  water,  and  very  gently  pass  it  between  the  fore- 
finger and  thumb  nails,  holding  it  vertically  to  remove  any  particles 
of  moisture  or  dirt.  Stretch  the  web  carefully  over  two  of  the  op- 
posite cuts  in  the  reticule  frame.  Fasten  one  end  by  a  drop  of  the 
shellac,  dropped  gently  from  a  bit  of  pointed  wood  or  the  blade 
of  a  penknife.  Wait  a  moment,  for  this  drop  of  shellac  to  harden. 
See  that  the  web  is  stretched  tight  across  the  frame,  and  apply 
another  drop  of  the  shellac  to  the  opposite  cut  with  its  enclosed 
web.  Wait  several  minutes  before  cutting  off  the  two  ends  of  the 
web,  and  then  proceed  in  the  same  manner  with  the  web  which  is  to 
be  placed  at  right  angles  to  this  one. 

One  of  the  best  spider-webs  for  this  purpose  is  obtained  from 
the  cocoons  of  a  species  of  spider  found  in  Michigan.  These  threads 
are  almost  opaque,  and  not  apt  to  relax  their  tightness  if  properly 
placed  on  the  diaphragm,  and  as  they  retain  their  elasticity,  they  are 
preferable  to  platinum  wires,  which  have  a  tendency  to  break,  owing 


36  COAL  MINE  SURVEYING 

to  their  great  brittleness.  The  best  spider- threads  are  those  of  which 
the  spider  makes  its  nest.  These  nests  are  yellowish-brown  balls, 
which  may  be  found  hanging  on  shrubs,  etc.,  in  the  late  fall  or  early 
winter.  The  nest  should  be  torn  open  and  the  eggs  removed;  if 
this  is  not  done,  the  young  spiders,  when  hatched,  will  eat  the 
threads.  The  fibers  next  to  the  eggs  are  to  be  preferred  on  account 
of  their  fineness  and  darker  color.  As  it  is  important  to  get 
the  proper  kind  of  spider-web,  the  following  letter  on  this  sub- 
ject from  Prof.  J.  B.  Davis,  University  of  Michigan,  Ann  Arbor, 
Mich.,  is  of  interest: 

"  The  species  of  spider  of  which  I  send  you  cocoons  is  not  difficult 
to  find  in  Ann  Arbor— Lat.  42°  26'  N—  as  far  as  my  experience  goes, 
and  is  numerous  on  Beaver  Island,  out  in  Lake  Michigan — about 
46°  N. — at  St.  James.  I  have  also  always  succeeded  in  hunting  it 
in  our  Michigan  woods,  in  places  of  concealment — under  bark  of  dead 
trees,  in  cracks  and  holes,  about  old  stumps,  logs,  and  the  like.  It 
is  especially  partial  to  painted  woodwork.  It  roosts  high — the 
higher  the  gable  the  more  numerous  the  cocoons;  but  it  is  also  found 
on  fences  quite  numerously,  as  I  am  led  to  think  it  is  quiet  rather 
than  security  this  spider  seeks.  The  body  of  the  female  is  about 
three-fourths  of  an  inch  long,  and  nearly  half  an  inch  wide  across  the 
abdomen.  The  male  is  about  the  same  length,  but  far  slimmer. 
They  are  both  entirely  harmless.  I  never  knew  any  one  to  get 
bitten  by  either,  and  many  persons  in  my  observation  have  had  them 
freely  crawling  over  their  hands,  face  and  body.  They  may  be 
certainly  gently  handled  without  the  least  harm.  They  both  (male 
and  female)  bear  a  plain  escutcheon  design  on  the  back  of  the 
abdomen;  female  much  the  more  beautiful — in  browns.  Colors  all 
brown  and  yellowish-brown.  The  cocoon  is  a  snarl  of  webs,  and  is 
attached  under  ledges  of  window-sills,  cornices,  projections  of 
gables,  and  the  like  partly  sheltered  places.  The  color  of  the  threads 
you  have  is  of  a  light  corn-color,  distinctly  separating  it  from  the 
white  cotton-like  cocoons  so  common  everywhere.  The  threads  are 
silky,  not  like  cotton.  Of  late  years  I  keep  one  or  two  nice  cocoons 
where  they  can  be  reached.  You  know  one  can  wrap  them  in  a  bit 
of  paper  and  carry  them  in  the  pocket,  or  any  such  place,  and 
they  are  always  ready." 

ACCIDENTS  AND  PRECAUTIONS  IN  THE  USE  OF  INSTRUMENTS 

It  cannot  be  denied  that  instruments  frequently  meet  with  serious 
accidents  which,  with  a  little  care  on  the  part  of  the  operator,  could 


CARE  OF  INSTRUMENTS  37 

be  prevented.  It  certainly  does  not  betoken  proper  care  to  leave  it 
standing  unguarded  in  a  street,  road,  or  pasture,  or  in  close  vicinity 
to  blasting,  or  to  expose  it  unnecessarily  to  the  burning  rays  of  the 
sun,  or  to  dust,  dampness,  or  rain  at  any  time.  Such  carelessness 
must  inevitably  result  in  deterioration  of  the  accuracy  and  efficiency, 
not  to  speak  of  the  durability,  of  an  instrument.  It  should  be  borne 
in  mind  that  there  are  many  parts  of  an  instrument  which,  if  once 
impaired,  cannot  be  restored  to  their  original  efficiency. 

Tripods. — Legs  of  tripods,  if  fitting  too  loose  or  too  tight,  and  dull 
shoes  are  frequent  sources  of  falls,  and  loose  shoes  tend  to  make  an 
unsteady  instrument.  The  test  of  the  proper  degree  of  the  tightness 
of  the  legs  is  if  the  leg  is  raised  to  a  horizontal  position  and  left  free, 
it  should  gradually  sink  to  the  ground.  If  it  drops  abruptly  it  is 
too  loose;  if  it  does  not  sink  it  is  too  tight. 

Mounting  the  Instrument. — 'When  taking  an  instrument  from  its 
box,  it  is  not  immaterial  where  and  how  to  take  hold  of  it.  To  lift 
it  by  the  telescope,  circles,  standards,  or  wyes  is  improper,  and 
while  it  may  not  be  attended  at  once  with  any  serious  consequences, 
yet  it  may  sometimes  lead  to  some  permanent  injury,  and  it  cer- 
tainly is  aways  fraught  with  danger  to  the  permanency  of  the  adjust- 
ments. In  handling,  it  is  always  best  to  place  the  hand  beneath  the 
leveling  base. 

When  mounting  an  instrument  on  the  screw  of  its  tripod, 'or  screw- 
ing any  of  its  parts  together,  it  is  important  to  turn  the  part  in  the 
direction  of  unscrewing  until  it  is  perceived  by  a  slight  jar  that  the 
threads  have  come  to  the  point  where  they  enter;  the  motion  may 
then  be  reversed,  and  the  parts  screwed  together. 

To  secure  an  even  wear  of  tangent  and  micrometer  screws,  they 
should  be  used  equally  on  all  portions  of  their  lengths. 

Carrying  an  instrument  in  cold  weather  into  a  warm  room,  without 
the  protection  of  its  box  or  bag,  will  cause  a  sudden  exchange  of  air 
within  the  hollow  spaces,  and  carry  with  it  dust  and  other  substances 
through  the  minutest  openings.  The  vapor,  also,  that  will  thus  con- 
dense on  the  metal  surfaces,  if  it  were  not  protected,  will  have  a 
tendency  to  settle  a  film  on  exposed  graduations,  making  them 
indistinct  and  difficult  to  read. 

Protection  of  Lenses. — failure  to  protect  the  lenses  of  the  eye- 
piece and  object  glass  of  a  telescope,  when  not  in  actual  use,  from 
the  effects  of  moisture,  dust,  etc.,  by  the  covers  provided  for  them 
(eyepiece  lid  and  cap)  will  result  in  a  more  frequent  settling  of  a 
thin  film,  which,  like  the  fatty  substance  left  by  the  touch  of  the 


38  COAL  MINE  SURVEYING 

fingers,  greatly  impairs  the  clearness  of  vision.  That  the  too 
frequent  cleaning  of  the  lenses  must  in  the  course  of  time  be  detri- 
mental to  their  brilliant  polish,  and  lead  to  a  corresponding  loss  of 
transparency,  so  essential  to  the  proper  working  of  a  good  telescope, 
is  apparent.  Too  much  care  cannot  be  taken  to  guard  the  lenses,  and 
particularly  the  inner  surfaces  of  the  lenses  comprising  the  objective, 
against  any  film  that  may  settle  on  them.  The  ill  effects  of  such 
a  film  are  especially  noticeable  in  high-powered  telescopes  of  first- 
class  geodetic  and  astronomical  instruments.  In  short,  it  should  be 
remembered  that  the  slightest  film,  scratch,  or  dirt  will,  according 
to  their  nature  and  location,  impair  the  sight  through  a  telescope,  and 
often  render  it  unfit  for  accurate  work. 

Glass  Parts. — The  glass  covers  protecting  the  compass,  arc,  and 
verniers  from  exposure  need  very  careful  brushing  and  cleaning,  the 
same  as  the  lenses,  as  any  scratch  or  film  will  impair  their  transpar- 
ency. If  at  any  time  the  ground-glass  shades  should  lose  their  pure 
whiteness,  by  either  dirt  or  film,  and  will  not  act  as  illuminators  of 
the  verniers  and  graduation,  take  them  out  of  their  frames  and 
simply  wash  them  with  soap  and  water. 

The  Needle. — To  prevent  loss  of  magnetism  in  the  needle  of  instru- 
ments provided  with  a  compass:  when  storing  away,  allow  the  needle 
to  assume  magnetic  north  and  south;  then,  by  means  of  the  lifter, 
raise  it  from  the  center-point  against  the  glass  cover. 

If  an  instrument  has  met  with  a  fall,  bending  centers  and  plates, 
etc.,  it  should  not  be  revolved  any  more,  in  order  to  preserve  the 
graduations  from  still  further  injury,  but  recourse  should  be  had  at 
once  to  the  nearest  competent  maker. 

Instrument  Boxes.— If  the  box  or  tripod  should  have  become  wet, 
they  should  be  rubbed  dry,  and  the  varnish  should  be  renewed 
whenever  found  wanting.  Loose  or  detached  resting-blocks  in  the 
instrument-box,  or  any  looseness  of  the  instrument  in  them,  are 
very  detrimental  to  the  instrument  and  its  adjustments.  Cracks 
in  the  instrument-box,  the  absence  of  rubber  cushions  under  it, 
worn-out  straps  and  defective  buckles,  hinges,  locks,  and  hooks, 
should  never  be  tolerated,  as  the  remedy  is  so  easily  applied  by  any 
mechanic.  Such  defects  and  imperfections  are  known  to  lead  to 
injury  of  the  instrument. 

Storing  Instruments. — The  place  where  instruments  are  kept  or 
stored  away  should  be  thoroughly  dry  and  free  from  gases.  The 
placing  of  fused  chloride  of  calcium,  or  caustic  lime,  in  an  open 
vessel  in  the  instrument-box  is  to  be  recommended  where  there  is 


CARE  OF  INSTRUMENTS  39 

dampness;  and  if  the  presence  of  sulphureted  hydrogen  is  suspected, 
then,  cotton  saturated  with  vinegar  of  lead,  placed  in  the  box,  will 
prove  a  preventive  against  the  tarnishing  of  solid  silver  graduations. 

TRANSPORTATION  or  INSTRUMENTS 

During  the  progress  of  field  work  the  more  ordinary  and  portable 
transits  and  leveling  instruments,  etc.,  can  generally  be  carried  on 
their  tripods  for  ease  and  dispatch.  Nothing  in  the  way  of  precise 
instructions,  however,  as  to  the  best  method  of  carrying  an  instru- 
ment, whether  on  the  tripod,  in  the  arm  without  the  tripod,  placing 
the  hand  beneath  the  leveling  base,  or  in  the  box,  can  be  suggested 
here.  The  nature  of  the  ground,  the  surroundings,  the  size  and 
weight,  and  the  distance  to  be  traveled  over,  and  last  but  not  least 
the  fineness  of  the  instrument,  will  dictate  to  the  engineer  the  best 
means  of  conveying  it  from  point  to  point  in  order  to  protect  it  from 
injury,  and  its  adjustments  from  derangement. 

Carrying  an  instrument  on  its  tripod  without  slightly  clamping  its 
principal  motions,  will  wear  out  the  centers.  When  carrying  on 
its  tripod,  clamp  telescope  in  the  transit,  when  placed  on  a  line  with 
its  centers  and  in  the  level  when  hanging  down. 

Placing  in  the  Box.— When  carrying  an  instrument  in  the  box  it 
is  important  that  it  be  placed  therein  exactly  in  the  position  and 
manner  designated  by  the  maker.  Therefore,  upon  receiving  a  new 
instrument,  the  first  step  should  be  to  study  its  mode  of  packing, 
and  if  necessary  a  memorandum  should  be  made  for  future  guidance 
and  pasted  in  the  box.  This  will  save  time  and  vexation,  as  some  of 
the  boxes  for  field  instruments  must  necessarily  be  crowded  to  be 
light  and  portable. 

Before  placing  an  instrument  with  four  leveling  screws  in  its  box, 
the  foot-plate  should  be  made  parallel  to  the  instrument  proper,  and 
then  brought  to  a  firm  bearing  by  the  leveling  screws.  The  instru- 
ment must  also  be  well  screwed  to  the  slide-board,  if  one  is  provided. 
Having  put  the  instrument  in  the  box  in  such  a  position,  that  no  part 
of  it  will  touch  the  sides,  the  principal  motions  are  now  to  be  checked 
by  the  clamp  screws,  to  prevent  motion  and  striking  against  the  box. 
With  instruments  not  standing  erect  in  their  boxes,  but  which  are 
laid  on  their  sides  in  resting  places,  padded  with  cloth,  specially 
'provided  for  that  purpose,  their  principal  motions  must  not  be 
clamped  until  the  instrument  has  been  secured  in  a  complete  state 
of  repose  in  these  receptacles,  so  as  to  be  entirely  free  from  any 
strain.  Care  must  be  taken,  too,  that  all  of  the  detached  parts  of  an 


I 
40  COAL  MINE  SURVEYING 

instrument,  as  well  as  its  accessories,  are  properly  secured  to  their 
receptacles  before  shutting  the  box. 

Shipping. — When  shipping  an  instrument  over  a  long  distance  it 
is  commendable  to  fill  the  hollow  space  between  it  and  its  box  with 
small  soft  cushions  made  of  paper,  or  of  excelsior  or  shavings  wrapped 
in  soft  paper,  .taking  care  not  to  scratch  the  metal  surfaces,  nor  to 
bend  exposed  parts,  nor  to  press  against  any  adjusting  screws. 

For  greater  safety  in  transportation  by  express,  the  instrument- 
box  itself  should  always  be  packed  in  a  pine- wood  box  one  inch  larger 
all  around.  For  the  ordinary  size  of  field  instrument  the  packing- 
case  should  be  provided  with  a  strong  rope  handle,  which,  like  the 
strap  of  the  instrument  box,  should  pass  over  the  top  of  the  case  and 
through  holes  in  the  sides,  the  knots  being  within  the  case  and 
strongly  secured.  In  cases  where  the  gross  weight  of  the  entire 
package,  as  prepared  for  shipment  in  the  above  manner,  exceeds 
40  or  50  lb.,  then  two  men  should  handle  it,  and  two  strong  rope 
handles,  one  at  each  end  of  the  packing  case,  should  be  provided. 
In  order  to  check  jars  and  vibrations  while  en  route,  the  loose  space 
between  the  instrument  box  and  the  packing  case  is  to  be  filled  with 
dry  and  loose  shavings. 

The  cover  bearing  the  directions  should  always  be  screwed  on  and 
marked  in  large  black  letters.  The  upper  halves  of  the  four  sides  also 
should  have  "care"  and  "keep  dry"  marked  in  large  letters  on  them. 
These  precautions  are  indispensable  for  safe  conveyance  while  in  the 
hands  of  inexperienced  persons,  as  without  them  messengers  will 
often  carry  them  wrong  side  up. 

The  tripod  needs  packing  simply  in  a  close-fitting  box.  If  not 
placed  in  a  box,  it  often  happens  that  legs  or  shoes  are  broken  off 
while  en  route,  or  that  the  tripod  head  becomes  bent. 

Many  hundreds  of  instruments,  packed  as  explained  above,  have 
been  shipped,  travelling  thousands  of  miles,  over  rough  roads,  on 
stages  and  on  horseback;  and  the  instances  are  so  rare  where  one 
has  become  injured  (and  then  only  through  gross  carelessness), 
that  this  mode  of  packing  must  be  regarded  as  the  only  proper  one 
for  conveying  instruments  of  precision  by  express  or  other  public 
carriers. 

Arriving  at  its  destination,  an  instrument  should  not  remain 
packed  up  with  cushions,  etc.,  any  longer  than  necessary.  The 
atmosphere  in  such  boxes  naturally  must  be  close  and  often  moist, 
and  consequently  has  a  tendency  to  produce  the  ill  effects  by 
moisture  mentioned  in  preceding  paragraphs. 


CHAPTER  IV 
.ADJUSTMENT  OF  INSTRUMENTS 

NOTE. — Abstracted  from  publications  of  the  C.  L.  Berger  Co. 

The  mechanical  and  optical  condition  of  instruments  used  in 
geodesy,  and  their  adjustments,  although  satisfactory  when  they 
leave  the  maker's  hand,  are  liable  to  become  disturbed  by  use.  It  is 
therefore  of  vital  importance  that  the  person  using  an  instrument 
should  be  perfectly  familiar  with  its  manipulations  and  adjustments. 
He  should  be  able  to  test  and  correct  the  adjustments  himself  at  any 
time,  in  order  to  save  trouble  and  expense,  as  well  as  to  possess  a 
thorough  knowledge  of  the  condition  of  the  instrument.  It  is  evi- 
dent that  if  the  character  of  an  instrument  is  not  properly  under- 
stood or  if  the  adjustments  are  considerably  out,  the  benefit  due  to 
superior  design  and  workmanship  may  be  entirely  lost.  Under 
these  circumstances  an  expensive  instrument  may  be  little  better 
than  one  of  lower  grade. 

In  the  best  types  of  modern  instruments  the  principal  parts  are 
so  arranged  that  they  can  be  adjusted  by  the  method  of  reversion. 
This  method  shows  an  existing  error  at  double  its  actual  amount, 
and  renders  its  correction  easy  by  taking  one-half  the  apparent 
error.  Thus  errors  of  eccentricity  and  inaccuracy  in  the  graduations 
are  readily  eliminated  by  reading  opposite  verniers  and  reversing 
the  vernier  plate  180°  on  the  vertical  center  and  taking  the  mean  of 
the  readings,  and  by  repeating  the  measurement  of  an  angle  by 
changing  the  position  of  the  limb  so  that  the  measurement  will 
come  on  different  parts  of  the  graduation.  The  striding  levels 
and  levels  mounted  on  a  metal  base  are  readily  tested  by  revers- 
ing their  position  end  for  end.  In  the  transit  plate  levels  the 
adjustment  is  assured  by  turning  the  vernier  plate  180°.  Errors  of 
the  line  of  collimation  are  detected  or  eliminated  by  reversing  the 
telescope  over  the  bearings,  or  through  the  standards,  as  the  case 
may  be.  In  short,  an  instrument,  the  important  parts  of  which 
are  not  capable  of  reversing  in  one  way  or  another,  cannot  be  ex- 
amined quickly  and  accurately. 


42  COAL  MINE  SURVEYING 

THE  TRANSIT 

If  the  instrument  is  out  of  adjustment  generally,  the  engineer 
will  find  it  profitable  to  follow  the  makers  in  not  completing  each 
single  adjustment  at  once,  but  rather  bring  the  whole  instrument 
to  a  nice  adjustment  by  repeating  the  whole  series. 

The  Bubbles. — After  setting  up,  bring  the  two  small  levels  each 
parallel  to  a  line  joining  two  of  the  opposing  leveling  screws.  Bring 
both  bubbles  to  the  center  of  the  level  tubes,  by  means  of  the  leveling 
screws.  Now  turn  the  instrument  180°  in  azimuth.  If  the  small 
levels  still  have  their  bubbles  in  the  center  of  their  tubes,  these  levels 
are  adjusted,  and  the  circles  are  respectively  as  nearly  horizontal  and 
vertical  as  the  maker  intended  them  to  be. 

If  the  bubbles,  however,  are  not  in  the  center  of  their  tubes,  then 
bring  them  half  way  back  by  means  of  the  leveling  screws,  and  the 
remaining  half  by  means  of  the  adjusting  screw  at  the  end  of  each 
of  the  level  tubes.  It  may  be  necessary  to  repeat  this  adjustment 
several  times,  but  when  made,  the  instrument  once  leveled  will  have 
its  small  levels  in  the  center  of  their  tubes  through  an  entire  rotation 
of  the  circle. 

To  Make  the  Adjustment  for  Parallax. — This  adjustment  common 
to  all  telescopes  used  in  surveying  instruments  is  that  of  bringing 
the  cross  hairs  to  a  sharp  focus,  at  the  same  time  with  the  object 
under  examination.  Point  the  telescope  to  the  sky,  and  move  the 
eyepiece  until  the  cross  hairs  are  sharp  and  distinct.  Since  the  eye 
itself  may  have  slightly  accommodated  itself  to  the  eyepiece,  test  the 
adjustment  by  looking  with  the  unaided  eye  at  some  distant  point, 
and  while  still  looking,  bring  the  eyepiece  of  the  telescope  before  the 
eye.  If  the  cross  hairs  are  sharp  at  the  first  glance,  the  adjustment 
is  made.  Now  focus  in  the  usual  manner  upon  any  object,  bringing 
the  cross  hairs  and  image  to  a  sharp  focus  by  the  rack-work  alone. 
A  point  should  remain  bisected  when  the  eye  is  moved  from  one 
side  of  the  eyepiece  to  the  other. 

To  make  the  Vertical  Cross  Wire  Perpendicular  to  the  Plane  of 
the  Horizontal  Axis. — Bisect  some  point  at  the  lower  edge  of  the 
field  of  view  of  the  telescope  by  means  of  the  tangent  screw  and  note 
whether  it  continues  bisected  by  this  cross  line  throughout  its  entire 
length  when  the  telescope  is  moved  in  altitude.  If  it  does  not,  and 
the  point  is  to  the  right  of  the  line  in  the  upper  part  of  the  field,  the 
adjustment  is  made  by  loosening  the  four  capstan-headed  screws, 
and  rotating  the  reticule  in  the  direction  of  a  left-handed  screw, 


ADJUSTMENT  OF  INSTRUMENTS  43 

until  the  point  remains  bisected  and  then  tighten  all  four  adjusting 
screws.  Again,  bisect  the  point  by  means  of  the  tangent  screw. 
It  should  now  remain  bisected  throughout  the  length  of  the  cross 
wire,  if  not,  this  operation  must  be  repeated. 

To  Adjust  the  Vertical  Wire.— When  that  is  to  be  alone  adjusted 
in  the  field,  it  is  usually  done  according  to  the  following  simple 
directions:  Level  up  the  instrument  approximately  and  select  two 
distant  points  in  opposite  directions,  preferably  in  the  same  hori- 
zontal plane,  such  that  the  vertical  cross  line  will  bisect  them  both 
when  the  telescope  is  pointed  upon  one,  and  then  the  telescope  is 
reversed  on  its  horizontal  axis.  After  bisecting  the  second  point 
selected,  revolve  the  instrument  in  azimuth  and  bisect  the  first 
point  again  by  means  of  the  tangent  screw.  Reverse  the  telescope 
on  its  horizontal  axis  again,  and  if  the  second  point  is  now  bisected 
the  adjustment  for  collimation  of  the  vertical  wire  is  correct.  If  it 
is  not  bisected,  move  the  vertical  wire  one-fourth  of  the  distance 
between  its  present  position  and  the  point  previously  bisected. 
Again  bisect  the  first  point  selected,  reverse  the  telescope  and  find 
a  new  point  precisely  in  the  new  line  of  sight  of  the  telescope;  these 
two  points  will  now  remain  bisected  when  the  instrument  is  pointed 
upon  them  in  the  manner  described  above,  if  the  adjustment  is 
correctly  made.  If  the  two  points  are  not  now  both  bisected,  the 
adjustment  must  be  repeated  until  this  be  the  case. 

To  Determine  Whether  the  Standards  are  of  the  Same  Height.— 
Suspend  a  plumb  bob  by  means  of  a  long  cord  from  a  height  say  of 
from  30  to  40  ft.  The  plumb  bob  may  swing  in  a  bucket 
of  water  to  keep  it  steady.  (Instead  of  a  plumb  line  the  reflection 
of  a  church  spire  or  edge  of  a  tall  building  or  any  other  convenient 
object  may  be  viewed  in  a  bucket  of  water.)  Level  the  instrument 
carefully,  and  point  upon  the  plumb  line  at  its  base.  If  the  plumb 
line  remains  bisected  throughout  its  entire  length  when  the  telescope 
is  moved  in  altitude,  and  then  the  telescope  reversed  and  again  made 
to  bisect  the  line  throughout  its  length  from  its  base  upward,  the 
adjustment  is  correct. 

Otherwise  make  the  adjustment  by  means  of  the  capstan-headed 
screw  directly  under  one  of  the  telescope  wyes.  Loosen  the  screws 
in  the  pivot  caps  and  turn  the  vertical  adjusting  screw  right  handed 
to  raise  the  wye  bearing  one-quarter  of  the  error  to  be  corrected. 
If  the  telescope's  axis  is  already  too  high,  the  vertical  adjusting  screw 
should  be  loosened  a  little  more  than  needed  and  then  by  the  screws 
of  the  pivot  cap  the  wye  bearing  should  be  lowered  until  it  just 


44  COAL  MINE  SURVEYING 

touches  the  vertical  adjusting  screw.  The  screws  of  the  pivot  cap 
must  now  again  be  loosened  and  the  wye  bearing  raised  by  a  right- 
hand  turn  of  the  vertical  adjusting  screw,  as  explained  above,  until 
the  telescope's  axis  is  in  the  correct  position.  If  this  is  not  done 
the  adjustable  bearing  is  likely  to  stick  and  not  rest  on  the  adjust- 
ing screw,  thus  causing  liability  to  derangement.  The  screws  in  the 
pivot  cap  should  then  be  turned  down  just  enough  to  prevent  loose- 
ness in  the  bearings. 

Instead  of  using  a  plumb  line  a  simpler  method  having  the  advan- 
tage of  not  requiring  the  instrument  to  be  leveled  up  carefully  is  as 
follows:  Set  up  the  instrument  as  near  as  may  be  convenient  to  a 
building,  say  about  20  ft.,  in  order  to  get  as  high  an  altitude  as  pos- 
sible. Level  up  only  approximately,  clamp  and  bisect  a  point  at  the 
base  by  the  tangent  screw.  Then  elevate  the  telescope  and  find  a 
well-defined  object  as  high  as  possible,  only  using  the  telescope's 
horizontal  axis.  Now  reverse  telescope  and  move  instrument  on 
its  vertical  center,  again  clamp,  and  bi-sect  the  point  at  the  base.  If 
when  the  telescope  is  elevated  it  bi-sects  the  high  object  selected  the 
adjustment  is  correct.  If  it  does  not,  proceed  as  described  in  the 
above  method. 

To  Adjust  the  Level  to  the  Line  of  Collimation  of  the  Horizontal 
Wire. — One  method  is  to  use  a  sheet  of  water,  or  where  that  is  not 
available,  two  stakes  which  are  driven  with  their  surfaces  in  the  same 
level  plane.  Level  up  the  transit  half-way  between  two  points  lying 
nearly  in  a  horizontal  line,  and  say  300  ft.  apart.  Drive  a  stake  at 
one  of  these  points,  place  the  rod  on  it  and  take  a  reading,  first 
bringing  the  bubble  to  the  middle  of  its  tube.  Point  the  telescope 
in  the  opposite  direction,  again  bring  the  bubble  to  the  middle  of  its 
tube,  and  drive  a  second  stake  at  the  second  point  selected  until  the 
rod  held  upon  the  second  stake  gives  the  same  reading  as  when 
held  upon  the  first  stake.  The  tops  of  these  two  stakes  now  lie  in 
the  same  level  line. 

Take  up  the  transit  and  set  it  outside  in  line,  as  near  as  it  can  be 
focussed  on  the  first  stake  and  level  up.  Now  read  the  rod  upon  the 
first  stake  with  the  bubble  in  the  center  and  then  upon  the  second. 
If  the  two  readings  agree,  and  the  bubble  is  in  the  middle  of  its  tube, 
the  adjustment  is  correct.  If  the  two  readings  do  not  agree,  then 
by  means  of  the  telescope's  tangent  screw  elevate  or  depress  the 
telescope  the  amount  required  until  the  horizontal  wire  reads  the 
same  on  the  distant  rod.  Next  refocus  on  the  near  rod,  take  a  read- 
ing, then  focus  on  the  distant  rod  and  see  if  the  readings  are  the  same, 


ADJUSTMENT  OF  INSTRUMENTS  45 

if  not,  by  means  of  the  tangent  screw  again  make  the  horizontal  wire 
read  the  same  as  on  the  near  rod.  Repeat  this  operation  until  both 
rods  read  the  same.  Now  with  the  horizontal  wire  bisecting  the 
distant  reading  make  the  adjustment  of  the  level  by  its  capstan- 
headed  nuts  until  the  bubble  is  in  the  middle  of  its  tube  when  the 
level  will  be  parallel  to  the  line  of  collimation. 

THE  LEVEL 

The  Telescope. — After  the  engineer  has  set  up  the  instrument  and 
adjusted  the  eyepiece  for  parallax,  as  described  under  the  engineer's 
transit,  the  horizontal  cross  wire  had  better  be  made  to  lie  in  the 
plane  of  the  azimuthal  rotation  of  the  instrument.  This  may  be 
accomplished  by  rotating  the  reticule,  after  loosening  the  capstan- 
headed  screws,  until  a  point  remains  bisected  throughout  the  length 
of  the  wire  when  the  telescope  is  moved  in  azimuth.  In  making 
this  adjustment,  the  level  tube  is  to  be  kept  directly  beneath  the 
telescope  tube.  When  made,  the  small  set-screw  attached  to  one  of 
the  wyes  may  be  set  so  that  by  simply  bringing  the  projecting  pin 
from  the  telescope  against  it,  the  cross  wires  will  be  respectively 
parallel  and  perpendicular  to  the  motion  of  the  telescope  in  azimuth. 

The  first  collimating  of  the  telescope  may  be  made  using  an  edge 
of  some  building,  or  any  profile  which  is  vertical.  Make  the  vertical 
cross  wire  tangent  to  any  such  profile,  and  then  turn  the  telescope 
halfway  round  in  its  wyes.  If  the  vertical  cross  wire  is  still  tangent 
to  the  edge  selected,  the  vertical  cross  wire  is  collimated. 

To  Make  the  Adjustment  of  the  Horizontal  Wire. — Select  some 
horizontal  line,  and  cause  the  horizontal  cross  wire  to  be  brought 
tangent  to  it.  Again  rotate  the  telescope  halfway  round  in  its 
wyes,  and  if  the  horizontal  cross  wire  is  still  tangent  to  the  edge 
selected,  the  horizontal  cross  wire  is  collimated. 

Having  adjusted  the  two  wires  separately  in  this  manner,  select 
some  well-defined  point  which  the  cross  wires  are  made  to  bisect. 
Now  rotate  the  telescope  halfway  round  in  its  wyes.  If  the  point  is 
still  bisected,  the  telescope  is  collimated.  A  very  excellent  mark 
to  use  is  the  intersection  of  the  cross  wires  of  a  transit  instrument 
using  same  as  a  collimator. 

To  Center  the  Eyepiece. — This  is  done  by  moving  the  opposite 
screws  in  the  same  direction  until  a  distant  object  under  observation 
is  without  the  appearance  of  a  rise  or  fall  throughout  an  entire  rota- 
tion of  the  telescope  in  its  wyes.  The  telescope  is  now  adjusted. 


46  COAL  MINE  SURVEYING 

To  Adjust  the  Spirit  Level  to  the  Telescope. — Bring  the  level  bar 
over  two  of  the  leveling  screws,  focus  the  telescope  upon  some  object 
about  300  ft.  distant,  and  put  on  the  sun-shade.  These  precautions 
are  necessary  to  a  nice  adjustment  of  the  level  tube.  Throw  open 
the  two  arms  which  hold  the  telescope  down  in  its  wyes,  and  care- 
fully level  the  instrument  over  the  two  level  screws  parallel  to  the 
telescope.  Lift  the  telescope  out  of  its  wyes,  turn  it  end  for  end  and 
carefully  replace  it.  If  the  level  tube  is  adjusted,  the  level  will  indi- 
cate the  same  reading  as  before.  If  it  does  not,  correct  half  the  devia- 
tion by  the  two  leveling  screws  and  the  remainder  by  moving  the  level 
tube  vertically  by  means  of  the  two  adjusting  nuts  which  secure 
the  level  tube  to  the  telescope  tube  at  its  eyepiece  end.  Loosen  the 
upper  nut  with  an  adjusting  pin,  and  then  raise  or  lower  the  lower 
nut  as  the  case  requires,  and  finally  clamp  that  end  of  the  level  tube 
by  bringing  home  the  upper  nut.  This  adjustment  may  require 
several  repetitions  before  it  is  perfect. 

To  Make  the  Lateral  Adjustment  of  the  Spirit  Level.— The  level 
is  now  to  be  adjusted  so  that  its  axis  may  be  parallel  to  the  axis  of 
the  telescope.  Rotate  the  telescope  about  20°  in  its  wyes,  and  note 
whether  the  level  bubble  has  the  same  reading  as  when  the  bubble' 
was  under  the  telescope.  If  it  has,  this  adjustment  is  made.  If  it 
has  not  the  same  reading,  move  the  end  of  the  level  tube  nearest  the 
object  glass  in  a  horizontal  direction,  when  the  telescope  is  in  its 
proper  position,  by  means  of  the  two  small  horizontal  capstan- 
headed  screws  which  secure  that  end  of  the  level  to  the  telescope 
tube.  If  the  level  bubble  goes  to  the  object-glass  end  when  that 
end  is  to  the  engineer's  right  hand,  upon  rotating  the  telescope  level 
toward  him,  then  these  screws  are  to  be  turned  in  the  direction  of  a 
left-handed  screw,  as  the  engineer  sees  them,  and  vice  versa.  This 
accomplished  the  vertical  adjustment  of  the  spirit  level  for  parallel- 
ism with  the  line  of  collimation  of  the  horizontal  wire  must  now 
again  be  verified. 

To  Make  the  Adjustment  of  the  Level  Bar. — Level  the  instrument 
carefully  over  two  of  its  leveling  screws,  the  other  two  being  set  as 
nearly  level  as  may  be;  turn  the  instrument  180°  in  azimuth,  and  if 
the  level  indicates  the  same  inclination,  the  level  bar  is  adjusted.  If 
the  level  bubble  indicates  a  change  of  inclination  of  the  telescope  in 
turning  180°,  correct  half  the  amount  of  the  change  by  the  two  level 
screws,  and  the  remainder  by  the  two  capstan-headed  nuts  at  the 
end  of  the  level  bar.  Turn  both  nuts  in  the  same  direction,  an  equal 
part  of  a  revolution,  starting  that  nut  first  which  is  in  the  direction 


ADJUSTMENT  OF  INSTRUMENTS  47 

of  the  desired  movement  of  the  level  bar.  Many  engineers  consider 
this  adjustment  of  little  importance,  preferring  to  bring  the  level 
bubble  in  the  middle  of  its  tube  at  each  sight  by  means  of  the  leveling 
screws  alone,  rather  than  to  give  any  great  consideration  to  this 
adjustment,  should  it  require  to  be  made. 

To  Adjust  the  Horizontal  Wire  so  that  the  Line  of  Sight  will  be 
Parallel  to  the  Spirit  Level. — To  make  the  adjustment  with  the 
stakes,  set  up  the  level  halfway  between  two  points  lying  very 
nearly  in  a  horizontal  line,  and  say  300  ft.  apart.  Drive  a  stake  at 
one  of  these  points,  place  the  rod  on  it  and  take  a  reading,  first 
bringing  the  bubble  to  the  middle  of  its  tube.  Point  the  telescope 
in  the  opposite  direction,  again  bring  the  bubble  to  the  middle  of  its 
tube,  and  drive  a  second  stake  at  the  second  point  selected  until  the 
rod  held  upon  the  second  stake  gives  the  same  reading  as  when  held 
upon  the  first  stake.  The  tops  of  these  two  stakes  now  lie  in  the 
same  level  line. 

Take  up  the  level  and  set  it  outside  in  line  as  near  as  it  can  be 
focussed  on  the  first  stake  and  level  up.  Now  read  the  rod  upon  the 
first  stake,  and  then  upon  the  second.  If  the  two  readings  agree, 
and  the  bubble  is  in  the  middle  of  its  tube,  the  collimation  is  correct. 
If  the  two  readings  do  not  agree,  change  the  horizontal  wire  to  read 
the  same  on  the  distant  rod  by  means  of  the  capstan-headed  screws 
near  the  eyepiece  in  the  inverting  telescope  and  furthest  from  the 
eyepiece  in  the  erecting  telescope.  Refocus  on  the  nearest  rod, 
take  a  reading,  then  focus  on  the  distant  rod  and  again,  by  means  of 
the  capstan-headed  adjusting  screws,  make  the  horizontal  wire  read 
the  same.  Repeat  this  operation  until  both  rods  read  the  same, 
with  the  bubble  in  the  middle  of  its  tube. 


CHAPTER  V 
ORGANIZING  AND  EQUIPPING  THE  FIELD  PARTY 

The  mine  surveying  party  varies  widely  according  to  the  practice 
of  the  different  companies  and  in  different  parts  of  the  country. 
The  party  may  be  made  up  of  anywhere  from  two  to  four  or  five  men. 
Occasionally,  the  engineer  is  called  upon  to  do  general  surveys  with 
only  one  assistant,  but  this  is  one  of  the  most  flagrant  examples  of 
lack  of  economy  that  it  is  possible  to  conceive.  However,  the  two- 
man  party  is  not  without  its  uses,  as  for  instance,  in  doing  rough 
room  sighting  and  in  leveling;  two  men  are  all  that  can  ordinarily 
work  to  advantage  on  such  work  as  this,  and  it  is  unnecessary  to 
provide  more. 

The  three-man  party  makes  a  well-balanced  and  efficient  corps. 
For  general  surveying  it  is  the  most  economical  and  no  doubt  the 
one  in  most  general  use.  It  consists  of  the  transit  man,  and  two 
chainmen  or  backsight  and  foresight  as  they  are  sometimes  called. 
With  a  rapid  instrumentman  in  charge,  there  will  be  little  delay  in 
such  a  party  as  this,  each  man  having  practically  about  all  he  can 
do  to  keep  up  his  end  of  the  work. 

Occasionally,  where  greater  speed  is  necessary,  as  well  as  more 
detail,  a  four-  or  even  five-man  party  is  used.  Such  a  corps,  how- 
ever, is  usually  split  into  two  divisions.  Thus,  one  section,  including 
the  chief  of  the  party,  takes  the  lead,  establishing  the  stations,  meas- 
uring the  distances,  and  taking  the  neccessary  side  notes.  The 
second  section,  which  is  the  transit  party,  follows,  turning  the  various 
angles,  and  also,  as  a  rule,  measuring  the  distances  as  a  check  on  the 
work  of  the  first  party. 

An  organization  such  as  this  has  obvious  advantages  over  the 
smaller  party.  Thus  the  transitman  can  confine  his  entire  attention 
to  his  instrument,  which  will  do  much  to  eliminate  the  inaccuracies 
in  this  work;  the  possibilities  for  error  are  greatly  enlarged  where 
the  transitman  is  obliged  to  divide  his  attention  between  his  instru- 
ment and  directing  the  work  of  the  other  members  of  the  party; 
particularly  is  this  so  where  his  assistants  are  untrained.  The  larger 
party  is  also  able  to  cover  the  work  in  greater  detail.  On  the  other 

48 


ORGANIZING  AND  EQUIPPING  THE  FIELD  PARTY          49 

hand,  a  corps  of  this  size  is  liable  to  be  unwieldly  and  is  a  rather 
cumbersome  proposition  to  handle  underground;  it  is  obviously 
less  economical  than  the  three-man  party. 

The  five-man  party  is  confined  more  particularly  to  the  anthracite 
fields,  where  the  conditions  vary  widely  from  the  ordinary  bituminous 
practice.  The  methods  of  the  Lehigh  Valley  Coal  Co.  were  de- 
scribed in  the  Engineering  and  Mining  Journal  as  follows: 

A  mine-surveying  corps  is  generally  composed  of  five  men,  back- 
sight, foresight,  second-noteman,  first-noteman  and  transitman,  the 
latter  being  in  charge  of  the  party.  The  backsight  carries  a  single 
rod,  transit  plumb-bob,  an  extra  supply  of  bob  cord  and  a  two- 
quart  can  of  oil.  His  duties  are  to  help  the  second-noteman  orient 
or  "set  up"  the  transit  over  the  "spot"  established  by  suspending 
the  bob  from  the  station.  This  is  the  temporary  or  trial  set-up. 
The  backsight  then  inserts  his  rod  into  the  station  hole  suspending 
the  bob  over  the  transit-head  point  on  the  telescope,  while  the  transit- 
man by  means  of  the  shifting  plate  permanently  orients  the  transit. 

The  foresight  carries  a  single  rod  with  a  cast-iron  bob,  paint  can 
and  brush.  He  selects  the  most  advantageous  position  for  each 
station,  both  as  to  extension  and  good  roof,  sounding  the  top  rock 
with  his  rod  or  T-drill.  It  is  also  the  foresight's  duty  to  help  the 
second-noteman  measure  roof  distance  and  record  seam  sections, 
besides  giving  sight  to  height  of  instrument  for  vertical  angle,  and 
painting  the  station  number  on  the  roof  near  the  station. 

The  second-noteman  is  provided  with  an  8-  or  loft,  measuring 
pole  which  he  uses  to  estimate  the  offsets  at  all  ribs  and  intermediate 
points.  The  first-noteman  acts  as  assistant  transitman  and  records 
such  notes,  other  than  transit  observations  as  are  required.  While 
the  backsight  and  second-noteman  are  setting  up  the  instrument, 
the  transitman  looks  up  his  references  and  backsight  course,  and 
records  at  least  some  of  the  notes  for  the  section  connected  with  the 
previous  survey. 

Chief  of  the  Party. — By  this  term  is  meant  the  man  in  charge  of  a 
four-  or  five-man  corps.  The  first  qualification  of  such  a  man  is  a 
broad  underground  engineering  experience.  It  is  essential  that  he 
be  thoroughly  familiar  with  idiosyncrasies  of  the  underground 
workings,  so  that  he  will  be  able  to  judge  promptly  and  accurately 
the  best  method  of  procedure  in  the  face  of  the  unprecedented 
obstacles  which  are  always  arising  on  mine  work. 

The  chief  of  the  party  leads  the  way,  accompanied  by  one  or  two 
chainmen.  It  is  preferable  that  one  of  these  latter  be  a  workman 
4 


5o  COAL  MINE  SURVEYING 

in  that  particular  mine  or  section  of  the  mine,  so  that  he  will  be 
perfectly  familiar  with  the  workings  and  able  to  give  the  chief  any 
information  he  may  require.  The  chief  proceeds  with  his  party, 
establishing  the  stations  at  the  most  advantageous  points,  to  insure 
the  greatest  rapidity  in  covering  the  ground.  This  party  also  usually 
measures  the  distances  between  stations  and  takes  the  side  notes. 
It  is  possible  to  exercise  a  great  deal  of  ingenuity  in  the  location  of 
the  stations  so  as  to  accelerate  the  work,  as  for  instance,  it  is  clear 
that  these  should  be  as  far  apart  as  practicable,  so  as  to  insure  the 
minimum  number  of  setups  of  the  instrument. 

Transitman. — It  is  becoming  difficult  in  modern  times  to  obtain 
good,  efficient  and  reliable  men  on  instrument  work  in  the  mines. 
Underground  work  is  naturally  not  attractive  to  the  average  young 
engineer,  although  when  once  broken  in  he  usually  prefers  to  be 
inside  during  extremes  of  weather,  either  hot  or  cold.  However, 
the  work  is  more  or  less  underpaid,  considering  the  qualifications 
demanded,  and  as  a  rule  the  engineer  scarcely  becomes  proficient  on 
his  work  till  he  is  either  advanced  into  the  operating  department  or 
branches  out  into  some  other  line. 

It  is  desirable,  but  by  no  means  necessary,  that  the  transitman 
be  a  college  graduate,  although,  as  a  matter  of  fact,  the  practical  man 
who  has  fought  his  way  up  by  hard  knocks  will  have  a  certain  ad- 
vantage around  the  mines.  The  most  essential  qualifications  of  a 
good  instrumentman  are  experience  and  an  unlimited  patience. 
While  it  is  a  comparatively  simple  matter  to  learn  how  to  run  a 
transit  and  read  the  vernier,  no  one  can  become  really  proficient  and 
accurate  without  long  experience.  In  fact,  it  might  be  said  that 
there  is  a  great  deal  of  intuition  about  running  an  instrument  on 
mine  work.  The  experienced  transitman  is  instinctively  and  per- 
haps unknowingly  applying  continual  little  checks  and  tests  to  his 
instrument,  his  work  and  himself,  that  perhaps  while  unimportant 
in  themselves,  when  all  combined,  they  determine  the  real  character 
of  the  man's  work. 

As  his  name  implies,  the  transitman  handles  the  instrument  work 
of  the  party  exclusively.  He  carries  (or  at  least  should  carry)  his 
instrument,  and  is  responsible  for  its  condition.  Upon  him  and  the 
head  chainman  or  foresight  depends  the  entire  speed  of  the  party; 
if  the  roof  is  hard  so  that  the  foresightman  is  delayed  in  getting  in 
the  new  stations,  the  party  may  be  obliged  to  wait  on  him,  but  as  a 
rule,  unless  the  instrumentman  is  very  rapid,  the  reverse  is  true. 

In  the  five-man  party  the  transitman  is  relieved  of  a  great  deal  of 


ORGANIZING  AND  EQUIPPING  THE  FIELD  PARTY  51 

responsibility  and  is  thus  able  to  give  the  actual  instrument  work 
closer  attention.  But  in  the  three-man  party  the  necessity  of  direct- 
ing the  location  of  the  new  stations,  assisting  with,  and  in  fact  quite 
often  doing  the  chaining  himself,  as  well  as  taking  the  side  notes 
and  providing  ways  and  means  for  overcoming  an  unending  series  of 
unexpected  contingencies,  devolves  on  the  instrumentman.  They 
are  apt  to  have  a  serious  effect  upon  the  accuracy  of  his  work, 
especially  if  he  is  of  an  irascible  disposition. 

Chainman  and  Backsight. — The  embryo  transitman  must  always 
serve  an  apprenticeship  on  chain  and  sight  work.  He  may  be  either 
an  ambitious  young  chap  around  the  mine,  or  a  recent  graduate 
from  a  mining  school.  The  head  chainman  on  the  three-man  party 
should  be  an  active  intelligent  man  and  one  who  can  be  relied  upon 
to  hold  the  tape  accurately.  The  brunt  of  the  work  usually  falls 
upon  him.  He  must  go  ahead  establishing  the  new  stations,  raising 
intervening  curtains,  and  exercise  care  and  judgment  in  selecting  the 
location  of  the  stations.  Unless  he  works  rapidly  the  transitman 
will  have  the  backsight  taken  before  the  foresight  is  ready.  The 
head  chain  is  favored,  however,  in  many  little  ways,  such  as 
being  "permitted"  to  carry  the  instrument  occasionally,  and  also 
practise  in  setting  it  up  during  intervals  when  the  party  is  waiting 
the  passage  of  a  trip  or  some  other  contingency.  The  rear  chainman 
or  backsight's  duty  consists  only  of  following  the  party  and  giving 
an  occasional  backsight. 


CHAPTER  VI 
ENTRY  SURVEYING 

Picking  up  the  Starting  Stations. — Having  the  party  finally 
organized  they  proceed  to  the  entry  in  the  mine  where  their  work 
is  to  start.  The  transitman  or  chief  of  the  party  refers  to  the  notes 
of  the  last  survey  in  this  entry  which  may  be  either  in  the  current 
book  for  that  mine,  or  the  one  preceding.  The  party  then  proceeds 
in  the  entry  to  the  last  two  stations  of  the  previous  survey,  these 
being  located  by  their  proximity  to  certain  rooms,  cross  cuts,  or  side 
entries,  as  shown  in  the  side  notes  of  the  previous  survey.  Thus 
according  to  the  side  notes  the  next  to  last  station  may  be  18  ft 
outside  of  room  No.  32,  and  the  last  station  41  ft.  inside  of  the  second 
cross  cut  inside  of  room  No.  32.  The  stations  are  further  checked 
up  by  the  rights  and  lefts  according  to  the  side  notes.  Occasionally 
where  heavy  timbering  is  being  done  a  station  may  be  shifted  2 
or  3  ft.  from  its  original  location,  and  yet  to  all  appearances 
be  the  same.  Such  a  contingency  does  not  happen  often  without 
effecting  the  distance  between  stations,  so  that  checking  this  will 
usually  suffice  to  prove  that  they  are  intact.  In  exceptionally  bad 
ground,  however,  it  is  well  to  set  up  the  instrument  at  the  next  to 
last  station  and  turn  the  last  angle  of  the  previous  survey  to  insure 
complete  accuracy. 

Setting  up  the  Transit. — There  are  two  ways  of  setting  up  the 
instrument  on  mine  work,  either  directly  under  the  station  or  by 
plumbing  down  from  the  station  and  establishing  a  point  accurately 
below  it  and  then  setting  up  over  this  point.  When  using  this  latter 
method,  it  is  the  duty  of  the  foresight  to  set  the  point  on  the  floor 
while  he  is  giving  the  sight  on  his  station.  This  method  is  slower 
than  setting  up  directly  under  the  station  and  is  more  liable  to 
inaccuracies;  it  is  not  to  be  recommended. 

Setting  up  under  a  station  appears  to  the  beginner  as  an  exceed- 
ingly difficult,  if  not  hopeless,  task.  However,  a  little  experience 
with  this  method  soon  makes  a  man  proficient  at  it,  and,  if  he  expects 
to  follow  mining  work,  he  will  do  well  to  master  this  in  the  beginning. 
When  setting  up  over  a  point,  it  happens  quite  often  that  in  moving 
around  and  setting  the  transit  leg,  the  point  established  on  the  floor 
is  disturbed;  when  this  occurs,  it  is  necessary  to  pull  the  transit  up 

52 


ENTRY  SURVEYING  53 

and  reset  the  point.  Sometimes,  after  the  instrument  is  entirely 
set  up,  it  may  be  thought  that  the  point  has  been  moved  when  it 
really  has  not.  In  addition  to  this,  when  setting  up  under  the  sta- 
tion, the  instrumentman  has  both  the  point  and  the  bubbles  directly 
under  observation  all  the  time,  whereas,  in  setting  up  over,  he  must 
either  have  two  lights  or  take  the  one  out  of  his  cap  and  put  it  down 
to  the  plumb  bob  to  see  how  his  center  is. 

When  setting  up  under  the  station,  the  instrumentman  first  hangs 
his  plumb  bob  in  the  station,  having  the  point  at  about  the  height 
which  he  wishes  to  set  the  instrument  The  tripod  plate  is  then 
brought  as  nearly  horizontal  as  it  is  possible  to  judge  by  the  eye. 
In  doing  this,  it  will  be  found  that  the  instrument  as  a  whole  has 
been  moved  to  one  side  from  the  station.  This  is  where  the  difficulty 
arises  in  setting  up  under  the  station.  Leveling  up  always  displaces 
the  point  and,  vice  versa,  in  getting  under  the  point,  the  instrument 
is  thrown  out  of  level.  It  is,  therefore,  necessarily  a  cut-and-try 
procedure.  With  a  little  experience,  however,  a  man  soon  becomes 
remarkably  adept. 

In  setting  up  under  the  station,  it  is,  of  course,  necessary  that  the 
instrument  be  provided  with  a  point  on  the  telescope  exactly  over 
the  theoretical  center;  nearly  all  mining  transits  now  have  such  a 
point.  This  will,  of  course,  only  be  over  the  exact  center  when  both 
the  horizontal  plate  and  telescope  of  the  instrument  are  exactly 
level.  The  first  time  the  instrument  is  leveled  up,  the  telescope 
bubble  should  also  be  leveled  at  the  same  time,  and  then  clamped 
securely  in  place.  After  that,  it  is  only  necessary  to  level  with  the 
lower  bubbles.  When  the  instrument  has  been  brought  within 
about  half  an  inch  of  the  exact  point,  the  remainder  can  be  gained 
by  loosening  the  lower  level  screws  and  shifting  the  head  bodily. 

The  new  transitman  is  very  prone  to  waste  time  in  unnecessary 
accuracy  in  setting  up  the  instrument.  On  the  other  hand,  he  may 
be  lacking  in  accuracy  when  same  is  essential,  and  it  is  well  for  him 
to  investigate  just  what  refinement  is  necessary  along  this  line. 
Thus,  for  instance,  in  a  sight  100  tt.  long,  an  error  in  reading  of  i  min. 
means  a  difference  of  .029  ft.,  or  about  f  of  an  inch;  in  a  sight  50  ft. 
long,  i  min.  amounts  to  .0125  ft.,  or  about  A  of  an  inch.  With  a 
sight  25  ft.  in  length,  as  may  occasionally  occur  in  a  very  crooked 
entry,  or  in  taking  <a  sight  through  a  cross  cut  for  a  check,  i  min. 
amounts  to  only  .00625  ft.,  or  about  iV  of  an  inch.  It  is  therefore 
seen  that  the  shorter  the  sight,  the  greater  care  and  accuracy  must 
be  exercised  in  setting  up  the  instrument. 


54  COAL  MINE  SURVEYING 

Another  difficulty  which  often  troubles  the  beginner  on  instrument 
work  is  locating  the  cross  hairs  when  sighting  along  a  dark  entry. 
An  interesting  discussion  occurred  in  Coal  Age  on  this  subject, 
started  by  an  engineer  of  the  Great  Western  Coal  &  Coke  Co.,  who 
said  in  substance: 

In  discussing  practical  problems  in  mine  surveying  with  several 
engineers  in  this  district,  I  find  that  many  of  them  are  still  using 
sheets  of  white  paper  torn  from  note  books,  etc.,  to  reflect  the  light 
upon  the  plumb  bob  in  taking  "  sights"  below  ground.  This  method 
requires  the  illumination  of  the  cross  hairs  in  the  instrument. 

I  find  that  a  small  piece  of  tracing  cloth  about  6X9  in.,  held  2 
in.  behind  the  plumb  bob  or  string  and  illuminated  from  behind, 
will  show  good  results  in  sighting  up  to  300  ft.  (depending  on  power 
of  telescope),  and  no  illumination  of  cross  hairs  is  necessary.  They 
will  stand  out  clear  upon  the  illuminated  tracing  cloth.  The  cloth 
will  easily  last  one  or  two  days. 

To  which  an  Indiana  engineer,  speaking  of  the  same  method, 
says: 

I  first  used  this  practice  about  three  years  ago,  at  the  mines  of  the 
Rock  Island  Coal  Mining  Co.,  Hartshorne,  Okla.;  and  we  are  using 
the  same  method  at  the  present  time  in  surveying  the  mines  of  the 
Consolidated  Indiana  Coal  Co.,  Hymera,  Ind.  The  method  is  an 
old  European  practice  and  is  explained  in  some  European  textbooks 
on  mine  surveying. 

While  a  Tennessee  engineer  offers  the  following  improvement: 

Referring  to  the  illuminating  of  cross  hairs  in  a  surveying  telescope, 
I  would  suggest  the  following  improvement:  Obtain  an  embroidery 
ring  and  insert  a  sheet  of  tracing  cloth  between  the  rings  and  clamp 
it.  It  is  as  tight  as  a  drumhead  and  you  can  see  the  cross  hairs  and 
plumb-bob  string  with  no  illumination  other  than  a  lamp  held  behind 
the  screen. 

A  Hibbing,  Minn.,  engineer,  in  the  Engineering  and  Mining  Jour- 
nal, found:  that  tracing  cloth  held  behind  a  plumb-bob  string  works 
well  when  the  mine  is  dry,  but  if  there  is  any  water  dripping,  the  cloth 
soon  becomes  soaked  and  useless.  To  overcome  this  difficulty  I 
use  the  heavy-oiled  sheet  in  which  the  blueprint  paper  is  wrapped  by 
the  manufacturer.  I  find  that  the  water  has  little  effect  on  it  and 
that  it  lasts  much  longer  than  tracing  cloth  does  under  the  circum- 
stances. Moreover,  it  gives  just  as  good  results.  The  method  is 
especially  good  when  an  acetylene  lamp  is  used  behind  the  paper 
instead  of  a  candle. 


ENTRY  SURVEYING  55 

But  another  writer  in  the  Engineering  and  Mining  Journal  found 
even  a  better  method  for  use  with  the  acetylene  lamp  as  follows: 

The  time-honored  method  of  giving  a  sight  for  the  underground 
transit,  has  been  to  use  a  piece  of  tracing  cloth  held  between  the 
plumb-bob  cord  and  the  candle.  With  the  growing  use  of  the  acety- 
lene lamp  for  surveying,  this  method  is  not  wholly  satisfactory. 
The  hot,  jet  flame  of  the  lamp,  unless  closely  watched,  is  destructive 
of  the  tracing  cloth.  One  method  of  avoiding  the  difficulty  is  to  use 
the  hemispherical  reflector  with  its  surface  somewhat  dulled.  A 
skillful  helper  can  so  hold  a  lamp  thus  equipped  as  to  conceal  the  flame 
behind  the  bob  while  the  top  part  of  the  reflector  forms  a  white  back- 
ground for  the  head  of  the  bob  and  the  string.  It  is  difficult,  however, 
to  keep  the  reflector  clean  of 
soot  and  rust  on  the  one  hand 
and  not  too  dazzlingly  brilliant 
on  the  other.  Obviously,  a  de- 
vice designed  for  the  purpose 
of  giving  sights  would  be  more 
satisfactory,  and  such  a  device, 
Fig.  17,  is  now  offered  for  sale. 

It  is  constructed  of  sheet  iron 
2  in.  wide  and  7\  in.  long  bent 
to  an  elliptical  curve  so  that  all 
parts  are  well  illuminated,  and 
painted  with  a  white  enamel.  FIG.  17.  REFLECTOR  FOR  THE  ACETY- 
It  has  a  hole  in  the  center  LENE  LAMP. 

through  which  the  lamp  burner 

projects.  A  small  piece  riveted  to  the  back  consists  of  a  nipple 
to  slip  over  and  grip  the  lamp  burner  and  two  eyes,  to  which  is 
attached  a  strap  embracing  the  lamp  and  holding  the  reflector  on. 
As  seen,  the  reflector  has  its  long  dimension  horizontal  when  the  lamp 
is  upright.  This  permits  it  to  be  worn  in  the  cap  without  danger  of 
its  striking  the  back.  In  use,  the  lamp  is  tipped  on  its  side  and  the 
flame  being  concealed  by  the  bob  so  as  to  eliminate  all  glare,  the 
reflector  then  forms  a  white  strip,  showing  the  top  and  bottom  of  the 
bob  and  an  inch  or  so  of  cord.  The  lamp  in  this  horizontal  position 
will  burn  long  enough  to  give  a  sight.  If  desired,  the  reflector  can 
be  easily  detached  and  carried  in  the  pocket. 

Not  only  is  this  device  quicker  and  neater  than  the  tracing  cloth 
method,  but  it  also  gives  a  clearer  sight  and  can  be  manipulated  with 
one  hand.  The  only  precautions  necessary  are  to  hold  the  reflector 
plumb  and  to  keep  the  lamp  flame  concealed. 


56  COAL  MINE  SURVEYING 

Turning  Angles.— There  are  three  general  methods  of  turning 
angles,  only  two  of  which  are  considered  good  practice  in  mine  sur- 
veying. Occasionally,  where  some  railroad  man  has  been  placed  in 
charge  of  mine  surveys,  we  find  that  he  adheres  to  the  old  method  ot 
turning  simple  rights  and  lefts.  Thus,  he  sets  the  vernier  on  zero, 
reverses  the  telescope,  takes  his  backsight,  plunges  the  telescope,  and 
turns  either  a  right  or  left,  as  the  case  may  be.  This  method  has  the 
disadvantage  in  that  inaccuracies  may  occur  in  setting  the  instrument 
on  zero  each  time,  and  it  does  not  permit  of  carrying  the  azimuth  as 
the  work  proceeds.  It  has  also  been  found  that  a  great  deal  of  con- 
fusion often  occurs  as  to  whether  the  sight  is  a  "right"  or  a  "left." 

The  system  of  doubling  angles,  while  rather  slow,  is  very  accurate 
and  has  a  great  deal  to  recommend  it,  although  it  also  has  the  dis- 
advantage of  not  giving  the  azimuth  which  is  often  so  essential  to 
have  on  mine  work.  However,  azimuths  can  be  computed  as  the 
survey  is  run  along  if  necessary,  although  it  is  well  to  eliminate  all 
work  of  this  kind  with  the  field  party,  since  it  involves  unnecessary 
delays  and  possible  inaccuracies. 

In  doubling  angles,  the  vernier  is  set  on  zero,  the  backsight  taken, 
and  the  complete  horizontal  angle  turned  to  the  foresight.  The  angle 
is  then  read,  the  lower  clamp  loosened,  and  another  backsight  taken 
without  disturbing  the  vernier.  The  plates  are  then  loosened  and 
the  complete  horizontal  angle  turned  again  as  before  by  taking 
another  foresight.  Obviously,  the  reading  of  this  second  angle 
should  be  twice  that  of  the  first. 

Assuming  that  the  first  reading  was  174°  20',  the  second  would,  of 
course,  be  348°  40'.  Or,  if  the  first  reading  was  greater  than  180°, 
say  for  instance,  186°  25',  the  second  must  necessarily  exceed  360°  by 
twice  as  much  as  the  first  was  greater  than  180°,  or,  in  this  instance, 
12°  50'.  In  the  practical  use  of  this  method  it  is  often  found  that  a 
discrepancy  of  one  minute  occurs  in  the  two  readings,  so  it  is  generally 
advisable  to  turn  the  angle  three  times  in  order  to  determine  which 
way  this  one  minute  should  be  thrown,  as  it  is  not  customary  to 
carry  fractions  of  a  minute  in  ordinary  mine  work.  In  addition  to 
eliminating  inaccuracies  in  reading  angles,  this  method  also  insures 
against  any  poor  adjustment  of  the  instrument  for  collimation  which 
is  often  quite  an  item  with  some  instruments  that  are  found  around 
coal  mines. 

The  method  of  continuous  vernier  is  by  far  the  most  rapid  and 
accurate  system  when  in  the  hands  of  a  good  man  equipped  with  a 
well-adjusted  instrument.  It  also  has  a  peculiar  advantage  in  that 


ENTRY  SURVEYING  57 

it  is  impossible  for  accumulative  errors  to  creep  in  on  a  long  survey 
as  is  the  case  with  either  of  the  two  previously  described  methods. 
In  this  system  the  instrumentman,  instead  of  setting  the  vernier  on 
the  zero,  places  it  on  the  azimuth  of  the  two  stations  he  is  starting 
from.  Thus,  for  instance,  assume  the  azimuth  of  these  stations  to  be 
184°  12';  the  vernier  is  accordingly  set  on  this  angle  and  the  back- 
sight taken  with  the  telescope  reversed.  Turning  the  telescope 
over,  the  instrument  is  then  pointing  on  the  true  azimuth  and  the 
party  is  ready  to  proceed.  The  foresight  is  taken  and  the  instrument 
picked  up  and  moved  ahead  to  the  new  station  without  disturbing 
the  vernier,  the  angle,  of  course,  having  been  read  and  recorded. 
Setting  up  at  the  new  station,  the  backsight  is  taken  with  the  tele- 
scope reversed,  and  the  operation  repeated  as  before. 

It  will  thus  be  seen  that  the  entire  operation  of  the  survey  is  carried 
forward  without  it  being  necessary  for  the  instrumentman  to  be 
obliged  to  set  his  vernier,  with  the  single  exception  of  at  the  starting 
point.  By  the  other  methods,  where  the  instrument  is  set  at  zero, 
each  time  a  mistake  of  say  a  quarter  of  a  minute  in  each  setting,  or  in 
reading  the  final  angle,  would  make  possible  an  ultimate  error  of  five 
minutes  in  a  survey  of  twenty  setups.  Such  a  contingency  could  not 
occur  in  the  continuous  vernier  method.  Furthermore,  a  mistake  in 
reading  one  angle  would  not  affect  the  accuracy  of  the  succeeding 
ones.  In  using  this  method  the  principal  inaccuracy  to  guard  against 
is  the  chance  of  the  vernier  being  moved  when  the  change  is  made 
from  one  station  to  another  as  may  inadvertently  occur  when  setting 
up  or  passing  through  a  curtain  when  the  slow-motion  screw  is  liable 
to  be  moved. 

Straight  Line  Work. — A  great  deal  of  the  mine  work,  particularly  in 
flat  seams,  consists  simply  of  straight  line  work,  and  instead  of  running 
a  meander  survey,  the  instrument  is  used  only  for  projecting  straight 
lines  ahead  or  turning  occasional  angles  for  starting  off  new  entries. 
Work  of  this  character  varies  a  great  deal  in  the  accuracy  required. 
In  a  rope  haulageway  the  engineer  must  exercise  the  greatest  caution 
to  avoid  the  possibility  of  any  deviation  from  an  absolutely  straight 
line.  This  also  applies  to  the  more  important  main  headings. 

On  such  work  it  is  well  to  test  the  accuracy  of  the  instrument  con- 
tinually. When  setting  up  at  the  last  station,  the  line  should  be 
checked  by  taking  more  than  one  backsight,  particularly  where  there 
is  a  possibility  of  the  station  having  been  disturbed,  as  is  the  case 
where  the  roof  is  bad.  In  projecting  the  line  ahead,  it  is  also  well  to 
check  the  instrument.  When  one  point  has  been  set  ahead,  the 


58  COAL  MINE  SURVEYING 

telescope  should  be  reversed  and  plunged  in  the  opposite  direction, 
in  order  to  assure  that  there  is  no  inaccuracy  in  adjustment. 

It  is  not  well  to  try  to  attempt  to  set  line  sights  at  too  great  a  dis- 
tance; after  one  point  has  been  obtained  near  the  face  the  instru- 
mentman  should  move  up  there  and  set  two  stations  from  that  point 
for  the  use  of  the  workmen.  On  butt  entries,  which  are  relatively 
short,  such  a  refinement  in  lining  is  seldom  necessary. 

The  practice  of  the  Consolidation  Coal  Co.  in  setting  sights  or 
pointers  was  described  in  Coal  Age  as  follows: 

"  Pointers"  are  set  as  often  as  required  to  facilitate  good  alignment. 
These  "pointers"  consist  of  two  steel  spads,  shown  in  Fig.  18, 
placed  in  the  roof  in  wooden  plugs  driven  into  holes  drilled  for  that 


Spad 


FIG.    l8.      METHOD    OF   SETTING    "POINTERS,"    CONSOLIDATION   COAL   CO. 

purpose.  They  are  spaced  not  over  3  ft.  apart  in  order  that  the 
strings  hung  from  the  spads  may  be  easily  seen  when  one  light  is  used 
in  lining  over  them.  All  "pointers"  indicate  the  center  line  of  the 
heading. 

It  is  sometimes  the  practice  to  place  the  sights  close  to  one  rib 
which  compels  the  miners  to  drive  their  places  with  great  accuracy. 
Commenting  on  this  method  Coal  Age  says: 

The  question  of  whether  the  sight  line  should  be  located  in  the 
center  of  the  entry  or  at  the  rib  is  one  of  preference.  Most  entry 
drivers  prefer  the  sight  line  located  about  i  ft.  from  the  off  rib  of  the 
entry,  or  on  the  opposite  side  from  the  entry  cross  cuts.  When  the 
line  is  located  in  the  center  of  the  entry  the  sights  are  often  disturbed 
and  frequently  lost.  In  this  position  also,  it  is  more  difficult  to  keep 
the  entry  straight  than  when  the  line  of  sight  is  close  to  the  rib. 

To  which  an  engineer  of  the  Lehigh  Valley  Coal  Co.  added  the 
following: 

Referring  to  the  question  of  the  best  location  for  placing  sights 


ENTRY  SURVEYING  59 

when  driving  an  entry,  permit  me  to  say  that  I  believe  a  location 
i  ft.  off  the  rib  is  proper  in  pitching  or  chute  places,  where  the  line 
would  then  -be  over  the  man  way  and  readily  accessible  not  only  for 
the  foreman  or  miner,  in  lining  up  the  place,  but  also  for  the  engineer, 
who  may  be  called  on  to  extend  the  line  farther  up  the  breast. 

In  a  flat  seam  or  roadway,  however,  I  believe  that  the  line  should 
be  set  at  about  8  ft.  from  either  rib,  assuming  the  places  are  driven  16 
ft.  wide;  the  sight  line  would  then  be  in  the  center  of  the  entry.  My 
reasons  for  choosing  this  position  are:  First,  if  the  seam  is  dirty,  the 
refuse  must  be  gobbed  on  the  rib,  and  8  ft.  will  allow  the  sight  line 
to  clear  the  gob  easily.  Second,  if  the  sights  are  any  distance  from 
the  face,  the  brattice  extended  from  the  inside  cross  cut  toward  the 
face  is  liable  to  interfere  with  the  line  of  sight,  unless  it  is  kept  at  a 
sufficient  distance  from  the  rib.  In  this  case,  allowing  5  or  6  ft.  for 
the  width  of  the  brattice,  a  distance  of  8  ft.  will  give  a  good  clearance 
for  the  sight  line. 

Chaining. — The  term  chaining  is  rather  a  misnomer  as  applied  to 
mine  surveying.  The  chain  proper  is  a  relic  of  the  older  days,  and 
is  used  entirely  on  outside  work,  principally  railroad  surveying;  it 
is  entirely  out  of  place  in  the  mine.  Instead,  a  thin  flat  wire  tape, 
such  as  has  already  been  described,  usually  from  300  to  500  ft.  in 
length,  and  equipped  with  a  good  practical  reel,  capable  of  withstand- 
ing considerable  rough  handling,  is  used.  These  tapes  are  usually 
only  graduated  every  5  ft. ,  although  occasionally  some  are  found  3  ft. 
and  even  2  ft.  With  good  chainmen  5  ft.  is  sufficient. 

With  the  three-man  party,  when  the  foresight  goes  ahead  to  put  in 
his  new  station,  he  should  take  the  large  end  of  the  tape,  with  the 
reel,  ahead  with  him.'  After  the  angle  has  been  read  he  selects  the 
nearest  5-ft.  graduation,  which  he  holds  accurately  at  the  plumb-bob 
string  and  the  instrumentman  pulls  the  tape  taut,  taking  care  to 
see  that  it  is  free  its  entire  length,  and  catches  the  exact  distance  on 
the  tape  with  his  finger;  he  then  measures  to  the  nearest  5-ft.  mark 
with  a  small  self-winding  5-ft.  steel  tape  graduated  to  hundredths. 
The  head  chainman  calls  out  the  mark  he  is  holding,  and  the  transit- 
man  makes  the  necessary  addition  or  subtraction  and  records  the 
distance. 

Accurate  measuring  underground  is  rather  difficult,  particularly 
when  the  chainmen  are  not  experienced  or  are  unreliable.  It  is  neces- 
sary that  the  tape  be  stretched  very  tight  in  order  to  eliminate  the 
inaccuracy  due  to  sagging,  and,  since  it  is  usually  little  more  than  a 
thin  wire,  it  is  difficult  to  hold  it  accurately  on  the  point.  At  the 


60  COAL  MINE  SURVEYING 

end  having  the  handle,  this  can,  of  course,  be  gripped  firmly,  but  the 
man  at  the  other  end  must  depend  upon  his  hands  alone.  Some 
devices  have  been  put  on  the  market  for  overcoming  this,  with  more 
or  less  good  results,  and  half  a  turn  of  the  tape  around  the  waist  is  also 
of  great  assistance.  The  instrumentman  usually  prefers  to  read  the 
tape  himself  and  take  his  own  measurements  in  so  far  as  he  is  able  to 
do  so,  unless  he  has  perfectly  reliable  chainmen. 

All  measurements  must  be  taken  in  a  horizontal  plane,  and  where 
this  is  impossible,  as,  for  instance,  on  a  slope,  the  angle  of  inclination 
is  taken  with  the  transit  and  the  slope  distance  measured.  Thus,  the 
head  chainman  usually  sets  his  plumb  bob  at  the  height  he  wishes 
to  measure  and  gives  the  transitman  a  sight  at  the  top  of  it.  When 
the  measurement  is  taken,  he  then  holds  the  tape  at  the  same  point 
and  the  transitman  measures  to  the  instrument.  Having  given  the 
angle  of  inclination  and  the  slope  distance  the  horizontal  distance  is 
easily  computed. 

Some  companies  still  continue  to  measure  to  only  tenths  of  a  foot, 
but,  as  a  rule,  most  of  them  now  measure  to  hundredths.  This,  of 
course,  depends  a  great  deal  on  the  character  of  the  work.  In  a  small 
mine  of  comparatively  limited  area  and  life,  a  tenth  of  a  foot  is  suffi- 
ciently close  for  all  practical  purposes,  but,  with  a  large  colliery,  the 
accumulated  error  where  only  tenths  are  used,  might  become  so  large 
in  time  as  to  involve  serious  possibilities.  Furthermore,  when  it  is 
understood  by  the  men  of  the  party  that  the  accuracy  demanded 
is  so  fine,  it  has  a  beneficial  psychological  effect. 

•  Taking  Side  Notes. — The  measurement  having  been  completed, 
the  party  then  proceeds  to  take  the  side  notes,  that  is,  locate  all 
rooms,  crosscuts,  side  entries,  and  the  ribs  of  the  entry  along  which 
the  survey  is  being  made.  The  tape  is  first  stretched  accurately  on 
line,  the  zero  end  being  held  exactly  at  the  instrument.  One  of  the 
chainmen  proceeds  along  the  entry,  stopping  at  each  room,  cross- 
cut, or  any  other  feature  which  it  is  desired  to  locate,  and  reading 
the  tape  at  that  point.  He  calls  this  out  to  the  instrumentman, 
and  also  gives  the  right  and  left,  that  is,  the  distance  from  the  tape 
to  each  rib  at  that  point. 

Sometimes  these  rights  and  lefts  are  measured,  but  more  ordinarily 
they  are  simply  estimated  by  stepping;  this  depends  on  the  scale 
to  which  the  map  is  drawn  and  on  the  general  standard  which  it  is 
desired  to  maintain.  Where  a  map  is  drawn  to  a  scale  of  200  ft.  to 
the  inch,  it  is  not  possible  to  show  the  rights  and  lefts  along  an  entry 
with  sufficient  accuracy  to  justify  any  particular  pains  in  taking  them. 


ENTRY  SURVEYING  6 1 

Making  Checks. — As  the  survey  proceeds,  the  instrumentman 
should  utilize  every  possible  opportunity  for  testing  and  checking  the 
accuracy  of  his  work.  Also  it  is  well  to  make  all  measurements 
twice.  With  unreliable  chainmen  this  will  serve  to  make  them  more 
cautious  about  their  work.  It  is  also  well  to  take  a  needle  reading 
with  the  compass  every  few  stations;  of  course,  this  cannot  be 
accepted  as  the  basis  for  any  readjustment  in  the  instrument  work, 
due  to  the  possibility  of  the  needle  being  attracted,  but  it  will  often 
serve  to  pick  up  a  large  error  of  a  number  of  degrees. 

Where  surveys  are  being  carried  in  two  parallel  entries,  an  occa- 
sional check  should  be  made  through  the  last  crosscut,  say  about  every 
1000  ft.  This  does  not  require  much  time,  and  establishes  the  accu- 
racy of  the  work  beyond  allquestion.  It  is  also  well,  when  the  oppor- 
tunity is  afforded,  to  make  a  tie  survey  between  adjoining  sets  of 
entries  when  a  room  or  a  connection  of  some  kind  has  been  made. 
This  affords  a  supplementary  check  in  addition  to  that  obtained 
between  the  two  parallel  entries,  and  further  verifies  the  work.  A 
good  engineer  will  never  fail  to  accept  an  opportunity  to  check  up  his 
work. 

Kinds  of  Stations. — The  character  of  station  used  varies  generally 
according  to  roof  conditions.  The  station  most  commonly  used 
consists  of  an  ordinary  horseshoe  nail  flattened  out,  preferably  under 
a  steam  hammer,  and  with  a  hole  punched  about  the  center,  suffi- 
ciently large  to  permit  the  plumb  bob  string  to  be  inserted  without 
any  trouble.  Sometimes,  where  roof  conditions  are  particularly 
favorable,  that  is,  a  firm  tenaceous  shale,  these  horseshoe  nails,  or  as 
they  are  commonly  termed,  spads,  are  driven  directly  into  the  roof, 
a  small  hole,  about  f  in.  in  diameter,  being  drilled  for  their  accommo- 
dation by  means  of  a  small  jigger  drill. 

More  commonly,  however,  the  spads  are  inserted  in  wooden  plugs 
driven  in  the  roof.  Sometimes  these  plugs  are  round  and  about 
f  in.  in  diameter,  and  the  hole  in  which  they  are  placed  is  drilled  by 
means  of  an  ordinary  carpenter's  brace  equipped  with  a  special 
chisel  bit;  they  are  usually  drilled  from  i  to  2  in.  deep.  In  addition 
to  this,  some  districts  use  a  plug  about  2  in.  square,  made  usually  by 
cutting  a  2X4  in  two  pieces,  and  about  2  in.  long  with  a  sharp  but 
rather  abrupt  point.  Holes  for  these  plugs  are  made  with  an  ordinary 
miner's  pick,  and  the  principal  advantage  of  this  type  of  station 
is  that  the  party  is  not  obliged  to  be  overburdened  with  any  excess 
tools. 

A  writer  in  the  Engineering  and  Mining  Journal  describes  another 


COAL  MINE  SURVEYING 


type  of  station  and  drew  the  following  interesting  conclusions  on 
this  subject: 

For  a  number  of  years  the  standard  spad  for  underground  survey- 
ing was  a  horseshoe  nail  with  the  head  flattened  and  punched  or 
slotted  to  receive  the  plumb  bob  cord.  A  few  years  ago,  a  substitute 
was  proposed  consisting  of  a  wire  nail  with  a  triangular  file-cut  half- 
way through  the  rim  of  the  head,  the  nail  being 
inclined  when  driven,  so  as  to  bring  the  notch  on 
the  upper  edge  ,of  the  head.  Considerable 
experience  with  both  types  has  brought  out  the 
following  points: 

(1)  The  wire  nail  is  stronger  and  better  resists 
chance  blows  or  blasting,  if  set  near  the  face. 

(2)  It  is  cheaper  and  easier  to  make. 

(3)  It  makes  the  plumb  bob  easier  to  hang 
in  the  first  place,  but  slower  to  adjust  for  height. 

(4)  It  is  more   accurate.      Any  one  size  of 
plumb  bob  cord  can  occupy  but  one  position, 
whereas  in  a  hole  the  cord  can  climb  the  side 
and  be  off-center  by  a  small  fraction  of  an  inch. 

(5)  It  is  not  suitable  when  it  must  be  set  in  a 
position  difficult  to  reach.     In  such  cases,  it  is 
customary,  with  the  old-style  spad,  to  hang  a 
permanent  loop  of  cord  down  to  a  point  within 
easy  access.     This  is  not  possible  with  the  wire 
nail. 


FIG.    19. 


Spad  Station 
FIG.    20.     DETAILS    OF   THE   SPAD   STATION. 


The  unusual  physical  conditions  prevailing  in  the  anthracite 
regions  has  resulted  in  a  distinctive  type  of  station  being  adopted 
there.  The  practice  of  the  Lehigh  Valley  Coal  Co.,  as  described  in 
The  Engineering  and  Mining  Journal  several  years  ago,  was  as 
follows: 

Two  general  methods  of  putting  survey  stations  in  a  mine  have 


ENTRY  SURVEYING  63 

been  successfully  used;  the  first  is  the  "drill-hole"  station,  and  sec- 
ond, the  "  spad"  station.  The  drill-hole  station  is  made  by  means  of 
a  small  "T  "-shaped  drill  with  a  |-in.  bit;  the  latter  may  be  mounted 
at  the  bottom  end  of  the  single  rod,  Fig.  19.  The  spad  station, 
Fig.  20,  consists  of  a  spad  (either  steel,  copper  or  bronze)  driven 
into  a  white-pine  plug  wedged  into  a  f-in.  drill  hole,  from  £  to  f  in. 
deep.  The  advantages  of  the  drill-hole  station  over  the  spad  station 
have  been  substantially  proved. 

Painting  Stations. — As  a  rule,  most  engineers  paint  a  circle  around 
the  stations  and  the  number  so  that  they  may  be  found  readily  and 
also  reduce  the  liability  getting  the  wrong  station.  Such  a  plan  has 
much  to  commend  it  but  it  also  has  the  disadvantage  of  attracting 
the  attention  of  mischievous  drivers  or  trapper  boys  who  will 
occasionally  destroy  the  station.  Should  it  be  desired  to  do  this, 
however,  a  writer  in  Coal  Age  offers  the  following  suggestion: 

The  indistinctness  with  which  engineers'  centers  and  bench  marks 
are  painted  on  the  roof  and  ribs  in  coal  mines,  would  largely  be  cor- 
rected were  slaked  lime  used  in  place  of  white  lead  as  a  pigment. 
The  lime  must  be  well  slake*d  or  it  \vill  burn  the  brush.  The  objec- 
tions to  lime  are  the  greater  weight  to  be  carried  and  the  change 
of  condition,  which  takes  place  as  it  begins  to  set,  often  changing  a 
too  liquid  mixture  to  one  that  is  too  stiff.  These  are  the  advantages: 
A  legible  mark  of  an  intense  white  whicti  will  last  for  years,  which  is 
scarcely  dimmed  by  smoke,  and  the  use  of  a  material  which  can 
be  found  in  almost  any  plant  and  which  costs  but  little.  Some 
people  prefer  to  add  salt  to  the  water  in  which  the  lime  is  slaked. 
Whiting  does  not  appear  to  be  preferable  to  lime.  If  possible,  it 
would  be  well  to  repaint  the  lime  marks  after  a  few  minutes'  time 
have  elapsed,  with  the  idea  of  making  them  more  strong  and  more 
prominent,  but  this  is  not  by  any  means  necessary. 

With  the  Consolidation  Coal  Co.  stations  are  marked  with  a  painted 
white  circle  around  the  spad  and  the  number  on  the  nearest  rib. 
Pointers  or  sigHts  are  designated  by  crosses,  the  center  of  the  cross 
being  the  spad. 

Numbering  Stations. — Although  many  engineers  believe  that  the 
same  station  numbers  may  be  used  in  different  entries  or  different 
parts  of  the  same  mine,  this  is  not  generally  considered  as  good  prac- 
tice. It  occasionally  happens  that  the  physical  conditions  in  some 
mines  are  such  that  there  is  little  probability  of  confusion,  but  num- 
bers are  cheap  and  to  remove  any  possibility  whatever  of  trouble  in 
this  connection,  it  is  just  as  well  to  have  no  two  stations  in  the  same 


64  COAL  MINE  SURVEYING 

mine  with  the  same  number.  The  following  method  of  the  Consoli- 
dation Coal  Co.,  as  described  in  Coal  Age,  is  an  excellent  system  to 
adopt. 

Each  mine  has  its  individual  set  of  station  numbers,  beginning  with 
one  and  running  up  consecutively,  except  in  room  work  where  the 
stations  are  more  or  less  temporary,  and  where  a  continuous  system 
wojild  cause  a  too  rapid  increase  of  station  numbers.  In  front  of  each 
room  is  set  a  station,  and  designated  in  the  notebook  by  a  fraction, 
the  numerator  of  which  indicates  the  room  number  on  that  heading 
and  the  denominator  the  station  number.  Thus  f  indicates  the  first 
station  placed  for  room  number  six,  |  indicates  the  third  station  in 
room  number  seven  of  that  particular  heading. 

Station  numbers  need  not  follow  consecutively  on  any  particular 
heading,  neither  are  they  spaced  at  particular  distances,  except  when 
turning  off  entries.  The  plus  system  of  numbering  stations  is  never 
used.  Thus  there  may  be  three  or  four  consecutive  numbers  on  one 
heading  and  the  next  higher  station  number  will  be  in  an  entry  in  an 
entirely  different  section  of  the  mine.  However,  as  each  mine  has  all 
its  survey  notes  recorded  in  a  book  for  that  particular  mine  alone,  it  is 
no  trouble  to  locate  the  last  station  number  used. 


CHAPTER  VII 
KEEPING  SURVEY  NOTES 

TRANSIT  NOTES 

The  accompanying  illustration,  Fig.  21,  shows  a  good  system  of 
keeping  field  notes  on  a  continuous  vernier  survey.  It  might  be 
well  to  state  here  that  there  are  almost  innumerable  ways  of  recording 
mine  survey  notes,  and  each  engineer  usually  works  out  a  system 
according  to  his  individual  ideas  on  the  subject.  However,  the 
method  here  shown  is  a  good  practical  system,  and  one  that  might  be 
adopted  by  all  engineers  with  few  variations. 

The  heading  "Survey  of  the  Thirteenth  South  Entry"  with  the 
personnel  of  the  party  and  date,  is  similar  to  that  explained  later  for 
leveling.  The  first  column  is  used  for  the  station  number;  that  is, 
the  two  stations  for  which  the  azimuth  and  distance  are  determined. 
The  second  column  is  for  the  azimuth,  and  the  third  for  the  bearing, 
this  latter  seldom  being  used  in  actual  practice  in  the  field,  although 
occasionally  it  is  convenient  to  note  the  needle  reading  in  order  to 
serve  as  a  check  on  the  azimuth.  The  fourth  column  is  used  for 
recording  the  vertical  angle  when  measuring  on  highly  inclined  seams, 
and  the  fifth  column  the  tape  reading  of  the  same.  The  last  column 
is  used  for  recording  the  horizontal  distance  which,  as  will  be  noted,  is 
that  most  commonly  given. 

Care  should  be  exercised  in  checking  and  noting  the  starting  sta- 
tions. Thus  it  is  well  to  note  at  the  beginning  of  each  survey,  as 
shown  in  the  figure,  the  station  from  which  the  work  is  started;  in 
this  case  it  is  shown  that  the  instrument  was  set  up  at  station  1361, 
and  that  the  backsight  was  taken  on  station  1360,  reference  to  which 
was  found  in  book  9,  page  64.  In  order  to  avoid  any  possibility  of 
error  or  misunderstanding,  the  azimuth  and  distance  of  the  starting 
stations  are  also  recorded  at  the  beginning  of  each  new  survey. 

As  will  be  observed  the  notes  for  the  survey  of  the  main  entry  are 
first  given,  after  which  a  similar  line  is  run  in  the  air  course.  It  will 
also  be  noted  that  the  survey  in  the  main  and  air  course  were  tied 
through  the  last  crosscut;  this  is  good  practice,  and  all  surveys 
should  be  balanced  up  in  this  way  at  certain  prescribed  intervals;  it 

6s 


66  COAL  MINE  SURVEYING 

does  not  necessarily  involve  much  time  as  the  next  to  last  station 
can  be  placed  in  front  of  this  crosscut,  and  an  extra  sight  taken 
through,  while  on  the  regular  entry  survey.  The  notes  for  this  work 
can  be  readily  followed  out,  and  it  will  be  observed  that  a  tie  of  zero 
degrees,  one  minute,  has  been  effected. 

Most  engineer's  field  books  have  the  right-hand  page  ruled  in  a 
somewhat  similar  way  to  cross-section  paper.  On  this  page  the  side- 
notes  corresponding  to  the  survey  notes  on  the  opposite  page  are 


FIG.    21.      SKETCH  SHOWING   A  TYPICAL  METHOD   OF  RECORDING   MINE 
SURVEY   NOTES. 

usually  kept.  Thus  it  will  be  noted  that  the  survey  of  the  main 
entry  started  at  station  1361,  which  is  inclosed  in  order  to  distin- 
guish it  from  the  other  figures  indicating  distances.  After  the  meas- 
urement has  been  taken,  the  tape  is  laid  on  the  bottom  and  the 
chainman  goes  along  and  picks  off  the  distances  to  all  the  points 
which  it  is  desired  to  locate.  In  this  instance,  we  find  that  he  comes 
to  the  first  rib  of  a  room  at  20  ft.,  the  second  at  28  ft.,  while  a  cross- 
cut is  found  at 4 1  ft.;  another  room  at  83,  and  so  on.  The  chainman 


KEEPING  SURVEY  NOTES  67 

at  the  same  time  calls  out  the  "rights  and  lefts";  that  is,  the  dis- 
tance from  the  chain  to  the  side  of  the  entry  on  each  side.  And  so 
the  party  proceeds  along  the  entry,  gathering  all  of  this  supplemen- 
tary data  which,  rilled  in  on  the  skeleton  of  the  mine  survey,  makes 
the  completed  mine  map. 

The  rooms  are  measured  and  sidenotes  taken  in  a  similar  fashion, 
it  being  customary  where  great  accuracy  is  desired  to  set  the  instru- 
ment up  in  front  of  each  room  and  take  a  rough  sight  to  the  face; 
this  is  hardly  necessary,  however,  in  a  well-systematized  mine  since 
these  rooms  should  all  be  driven  on  centers.  Where  the  rooms  are 
working,  that  is,  still  being  driven,  it  is  well  to  put  a  "  W"  at  the 
face,  as  noted  in  the  sketch,  Fig.  21,  and  an  "S"  where  they  have 
been  stopped,  as  will  be  noted  in  the  next  to  the  last  room  which  has 
evidently  been  stopped  for  some  reasons.  A  great  many  companies 
also  require  their  engineer  to  take  a  section  of  the  coal  at  the  face 
of  the  entry  at  the  time  of  the  survey.  A  convenient  method  of 
sketching  such  a  section  is  also  shown.  Where  no  regular  maps  are 
kept  for  recording  sections  of  the  seam  such  information  may  prove 
of  some  value  at  times,  and  it  is  well  for  the  engineer  to  acquire  the 
habit  of  doing  this. 

The  methods  of  keeping  notes  in  the  anthracite  regions  vary 
considerably  from  that  in  the  bituminous  mines.  The  practice  of 
the  Lehigh  Valley  Coal  Co.,  which  may  be  accepted  as  typical  of 
the  hard-coal  districts  in  general,  was  described  in  the  Engineering 
and  Mining  Journal  as  follows: 

A  special  volume  is  kept  for  transit  readings,  while  "offset"  notes 
are  recorded  in  a  separate  book.  Fig.  22  represents  a  page  of  notes 
as  recorded  by  the  first-noteman.  The  date,  seam  and  organiza- 
tion is  noted  at  the  foot  of  the  page  and  all  observations  recorded 
reading  from  bottom  to  top. 

The  figures  along  the  left  margin  [(a)  Fig.  22],  indicate  the  plusses 
on  the  tape  with  o  at  station  450.  Figures  to  the  right  and  left 
of  the  dotted  line  (transit  line)  indicate  offsets  to  ribs.  The  meas- 
ured distance  between  stations  450  to  462  is  also  recorded.  At  the 
face,  station  462  +  10,  a  dip  of  10°  "in  and  to  the  left"  is  indicated; 
it  is  also  noted  that  section  60  (S-6o)  was  taken  on  the  right  rib  at 
face.  Offsets  are  taken  at  2o-ft.  intervals  and  are  called  out  to 
the  first-noteman  \vho  records  the  figures. 

Vein-section  measurements  are  noted  by  the  second-noteman 
and  verified  by  the  first-noteman,  as  is  also  the  dip,  direction  and 
strike  of  the  seam.  The  sections  are  recorded  in  the  back  of  the 


68  COAL  MINE  SURVEYING 

side-note  book,  writing  forward,  fully  referenced.  As  for  example, 
S-6o,  page  — ,  station  462+10.  A  section  of  seam  is  recorded  as 
follows: 


HARD  SANDSTONE  ROOF 


=  Benches  have  no  parting  0.4  <  Boney  coal 

2.7  =  Coal 
=  Benches  have  parting        3 .  i  =  Slate 

4.1  <  Coal 

4.9  =  Boney  coal  (streaked) 

7.9  <  Coal 


3-Q 
6-3 


Slate  or 
Bone. 


o'.4 


0.8 


1.6 


Total. . . 


•7-9 


10 

> 

16  \  7 

S6ff 

15 

Roll 

WIDMHl 

^" 

16— 

462 

83.90       ,24' 

ft'*  9 

463 

84.51 

8    \ 
10  '  13 

00 

14  \  8 

82 

I2\I2 

26' 

60 

II  \I2 

75 

& 

I3\IO 

42 

IS\IO 

70 

!j 

M 

10  \  12 

55 

10  \ 

53 

If 

12  \IO 

30 

\I2 

40 

10  \I2 

25 

8  'IS 

22 

12  \IO 

\ 

20 

10  \I2 

10 

ia\i2 

450 
Feb.l.li 

W8  ffoss  Vein 

6 

455 

ct) 

Perry]  &  Smith 

1 

a  b  COM.  AOS 

FIG.    22.      METHOD  OF  TAKING   SIDENOTES  IN  THE  ANTHRACITE  REGIONS. 


The  notes  in  (b~),  Fig.  22,  indicate  a  roll  or  fault  from  station 
455  +  53  to  +  88  along  left  rib.  Also  a  roll  in  the  face  dipping  16° 
and  in  a  diagonal  direction  from  right  to  left,  and  indicated  as  a 
down- throw. 

The  transit,  after  set-up  is  satisfactory,  is  sighted  to  the  backsight 
station,  and  then  the  new  station  is  observed  and  readings  recorded 
as  shown  in  the  accompanying  table: 


KEEPING  SURVEY  NOTES 


69 


Remarks 

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Horizontal 
distance 

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Difference 
in 
elevation 

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10           vo     to           vo      0            vo 

1 

oOOOOOOioOOM 
MHtoOOtoOOO 

70  COAL  MINE  SURVEYING 

As  will  be  noted  the  horizontal  angle  is  measured  in  azimuth,  with 
o°  as  south,  90°  =  west,  180°  =  north  and  270°  =  east. 

Station  450  is  the  set-up  station,  and  318  the  backsight  station. 
The  roof  distance,  in  order  to  obtain  new  datum  for  the  levels,  is 


3rdQuarttrly  Posting  -  Sept  20,1915 
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S/3° 

'[is  symbol  for  station                                                             Cow.  AM 

FIG.    23.      THE   CONSOLIDATION   COAL  CO.'S    METHOD    OF    KEEPING 
MINE-SURVEY   NOTES. 

measured  and  recorded.  In  the  third  column  the  magnetic  bearing 
of  the  line  connecting  the  two  stations  is  recorded;  this  is  worked  out 
during  spare  moments  in  the  mine  and  is  obtained  by  deducting 
33^°  55'  from  360°  =  s.  Readings  between  o  and  90°  are  direct 
courses;  between  90°  and  180°  are  deducted  from  180°;  between 


KEEPING  SURVEY  NOTES  71 

180  and  270°  deduct  180,  and  between  270°  and  360°  are  deducted 
from  360°.  The  differences  are  the  magnetic  courses  for  correspond- 
ing quadrants. 

In  the  fourth  column  for  stations  450-462  will  be  noted  ~T~~O — r 

This  indicates  that  the  vertical  angle  -f-  5°  15'  was  read  at  1.54 
ft.  above  H.  I.  (hight  of  instrument),  and  the  corrections  are  ap- 
plied in  the  office  work  for  calculating  the  correct  "difference  in 
elevation."  This  is  obtained  in  the  following  manner:  The  cosine 
of  5°  15'  times  the  tape  distance  83.90  equals  horizontal  distance 
83.55  ft->  and  the  sine  times  the  tape  distance  equals  7.68  ft.,  the 
vertical  distance,  but  as  the  angle  was  measured  1.54  ft.  above  H.  I. 
the  true  distance  in  elevation  is  7.68  —  1.54  =  6.14  ft.  Similarly 
a  correction  of  i  ft.  is  necessary  for  difference  in  elevation  for  sta- 
tions 401-464,  where  the  angle  was  measured  i  ft.  below  H.  I. 
The  sine  for  the  distance  and  angle  equals  5.73  ft.,  therefore  the 
true  difference  in  elevation  is  5.73  +  i.o  =  6.73  ft.  The  algebraic 


FIG.    24.      PLOTTED   NOTES   SHOWN   IN    FIG.    22. 

sign  for  these  corrections  represents  the  field  operation  and  is  the 
reverse  of  the  algebraic  correction. 

The  methods  of  recording  sidenotes  vary  widely  in  all  parts  of 
the  country  and  even  in  different  mines  in  the  same  district. 

The  Consolidation  Coal  Co.'s  method  of  recording  sidenotes  was 
described  in  Coal  Age  as  follows: 

In  locating  the  various  features  of  the  mine  work,  sidenotes  are 
taken  from  the  established  line  of  sight.  The  page  of  survey 
notes  shown  in  Fig.  23  demonstrates  this  more  clearly.  It  will  be 
noted  from  these  records  that  sights,  or  courses,  are  always  checked 
between  stations  in  order  to  make  sure  that  the  stations  still 
occupy  the  same  positions  they  did  when  they  were  set  and  the 
readings  taken;  also  to  catch  up  the  possibility  of  an  error  being 
made  when  the  stations  were  set.  The  accompanying  Fig.  24 
shows  the  entry  plotted  from  these  notes.  At  the  face  the  date 
of  the  survey  is  marked  to  show  the  position  the  heading  occupied 
at  that  time. 


7-1 


COAL  MINE  SURVEYING 
LEVEL  NOTES 


The  illustration,  Fig.  25,  shows  the  most  approved  method  of  keep- 
ing level  notes.  The  notes  are  divided  into  five  columns,  the  first 
being  for  the  station  numbers,  the  second  "B.  S."  for  the  back- 
sight, the  third  "H.  I."  the  hight  of  instrument,  the  fourth,  "F.  S." 
the  foresight,  and  the  fifth,  the  elevation. 


FIG.    25.      SKETCH  SHOWING  THE  CUSTOMARY  METHOD  OF  RECORDING   MINE  LEVEL 
NOTES. 

Let  it  be  assumed  that  instructions  have  been  issued  to  extend 
the  levels  in  the  i$th  South  entry  into  the  face,  no  special  data 
being  required,  just  the  general  elevations  incident  to  the  ordinary 
leveling  process.  The  engineer  refers  to  the  notes  of  the  last  levels 
and  finds  that  they  were  concluded  on  station  1517.  He  accordingly 
proceeds  in  the  entry,  locates  the  required  station  and,  after  assur- 
ing himself  in  every  possible  way  that  it  has  not  been  disturbed, 
he  is  ready  to  proceed  with  the  new  work. 

Before  starting  he  carefully  arranges  the  headings  in  his  notebook 


KEEPING  SURVEY  NOTES  73 

as  shown.  Too  much  importance  cannot  be  laid  on  the  necessity 
for  careful  references  and  cross  references  of  all  kinds.  First,  the 
"Levels  in  i5th  South"  is  of  course  essential  to  show  where  the  work 
is  performed.  Next  it  is  well  to  note  the  personnel  of  the  party  as 
"Evans,  instrument"  and  "Boggs,  rod."  One  of  the  chief  reasons 
for  giving  such  close  attention  to  this  matter  is  that  grave  questions 
may  at  times  hinge  on  the  accuracy  with  which  the  engineer's  work 
has  been  performed.  In  all  organizations  of  any  importance, 
there  are  always  men  who  establish  a  record  for  accuracy  and  con- 
scientious work,  and  it  is  to  be  regretted  that  there  are  some  just 
the  opposite.  Obviously,  therefore,  it  is  often  a  matter  of  great 
importance  who  the  work  was  done  by. 

After  setting  down  the  date,  the  engineer  then  proceeds  fto  note 
carefully  the  initial  or  starting  point  for  his  work,  which,  in  this 
case,  is  station  1517,  or  as  noted:  "Reversed  rod  on  Sta.  1517, 
see  bk.  9,  p.  1 24."  This  data  should  be  carefully  noted  for  the  reason 
that  should  subsequent  work  show  the  elevation  for  this  Sta.  1517 
to  be  in  error,  it  is,  of  course,  essential  that  this  be  known  in  order 
to  make  the  proper  corrections  in  the  succeeding  work.  The  ele- 
vation of  Sta.  1517  as  obtained  from  bk.  9,  p.  124  was  found  to  be 
1321.17,  and  this  is  accordingly  set  down  in  the  column  under  ele- 
vation. The  instrument  is  then  set  up  and  a  reversed  rod  reading 
taken  on  the  station. 

This  idea  of  establishing  "bench  marks"  on  the  mine  stations 
in  the  roof  varies  entirely  from  anything  practised  on  surface  work. 
The  reason  for  this  is  that  it  is  exceedingly  difficult  to  get  any  re- 
liable permanent  points  on  the  bottom  which  are  not  in  danger  of 
being  disturbed.  The  bottom  in  a  live  mine  is  more  or  less  con- 
stantly on  the  move,  due  to  varying  roof  loads,  and  is  also  liable  to  be 
taken  up  in  order  to  gain  more  head  room.  So  it  has  become  the  ac- 
cepted practice  to  establish  all  permanent  reference  pointsin  the  roof. 

While  it  is  customary  to  add  all  backsights  to  the  elevation  in 
order  to  obtain  the  height  of  instrument,  this  operation  is,  of  course, 
the  opposite  where  a  reversed  rod  reading  is  used,  the  backsight  being 
subtracted  from  the  elevation  instead  of  added.  With  the  reversed 
rod  reading  of  minus  1.13,  as  noted,  the  height  of  instrument  there- 
fore becomes  1320.04.  The  rodman  then  moves  as  far  ahead  as  the 
levelman  can  see  him,  which,  in  this  case,  is  at  Room  28.  For 
convenience,  and  as  a  possible  future  check,  he  holds  the  rod  on  the 
point  of  frog  of  this  room.  The  reading  at  this  point  is  found  to  be 
4.19,  which  indicates  that  the  new  station  is  that  distance  beneath 


74 


COAL  MINE  SURVEYING 


the  height  of  the  instrument,  and,  accordingly,  the  difference  be- 
tween the  two  gives  the  new  elevation  15.85.  In  keeping  level  notes, 
it  is  not  the  practice  to  carry  the  hundreds  of  feet  along  in  either 
the  height  of  instrument  or  elevation  column,  and  these  are  accord- 
ingly omitted  except  where  a  change  occurs  as  from  1300  to  1200 
as  is 'noted.  The  instrumentman  now  moves  his  instrument  up 
ahead  of  the  rod  as  far  as  he  can  see,  and,  taking  a  backsight,  ob- 
tains a  reading  of  0.53,  which  means  that  his  instrument  is  this 
amount  above  the  station  and  the  new  height  of  instrument  is, 
therefore,  16.38. 

And  so  the  operation  is  continued  until  we  come  to  the  face  of 
the  entry,  at  which  point  another  bench  mark  is  taken  on 
Sta.  1521  for  use  in  making  future  extensions.  As  in  the  previous 
case,  this  is  a  reversed  rod  reading,  so  that  instead  of  subtracting 
the  foresight,  as  we  ordinarily  do,  this  is  added. 

Now  we  arrive  at  the  method  of  balancing  the  leveling  notes. 
A  little  consideration  of  the  level  notes  will  show  the  reader  that  the 
difference  between  all  the  foresights  added  together,  and  all  the 
backsights,  must  give  the  difference  in  elevation  between  the  starting 
point  and  the  concluding  stations.  But  the  reversed  rod  readings 
introduce  complications  in  this  method;  however,  since  these  are 
confined  entirely  to  the  initial  and  concluding  stations  of  the  survey, 
we  may  disregard  them  and  confine  our  check  to  the  intervening 
stations.  Accordingly,  adding  up  the  backsights  and  omitting  those 
having  a  circle  (reversed  rod  readings),  we  get  2.72,  and  doing  the 
same  with  the  foresights,  we  obtain  25.64,  the  difference  between 
these  two  being  22.92.  Now,  subtracting  the  difference  between  the 
second  elevation  from  each  end.  that  is,  1315.85  and  1292.93,  we 
obtain  a  perfect  check,  22.92. 

Leveling  practice  of  the  Consolidation  Coal  Co.  was  described 
in  Coal  Age  as  follows: 

All  elevations  on  the  inside  are  run  by  precise  levels  starting  from 
benchmarks  on  the  outside  established  from  U.  S.  G.  S.  benchmarks. 
The  readings  are  taken  on  the  bottom  of  the  seam  every  100  ft. 
and  in  some  localities  closer,  depending  upon  the  irregularity 
of  the  coal.  Every  six  months  the  levels  are  advanced  to  the  breasts 
of  the  working  places,  no  elevations  being  taken  in  rooms  except 
for  special  purposes.  Bench  marks  are  also  established  ahead, 
usually  being  placed  on  a  station  or  pointer  in  the  roof  and  B.  M. 
marked  on  the  rib  opposite.  The  method  of  recording  inside  level 
notes  is  shown  in  Fig.  26. 


KEEPING  SURVEY  NOTES 


75 


CHAPTER  VIII 


SOME  PROBLEMS  IN  SURVEYING 

The  following  problems  have  been  selected  from  among  those 
submitted  to  Coal  Age  in  the  past  two  years;  they  represent  a  typical 
assortment  of  the  many  computations  that  the  colliery  engineer  is 
apt  to  encounter. 

Problem  i. — Given  the  data  shown  in  Fig.  27  and  it  is  desired 
to  know: 

(a)  The  length  of  a  proposed  incline  AB 

(b)  The  depth  of  a  vertical  shaft  EC 

(c)  The  distance  DC  from  the  foot  of  the  shaft  to  the  mouth  of 
the  drift  in  the  lower  seam. 

According  to  the  law  of 
sines  we  know  that:  The 
ratio  of  any  two  sides  of  a 
triangle  is  equal  to  the  ratio 
of  the  sines  of  the  opposite 


angles. 

In  the  triangle  ADB, 
designate  the  angles  by 
the  large  letters  A,  D  and 
B,  respectively;  then: 


Lower  Sean 


FIG.    27. 


By  the  law  of  sines 


A  =  20° 

D  =180—30  =150° 
B  =  30—20  =   10° 


AB 
AD 


sin  D 
sin  B 


sin    10 

But  since  the  sine  of  an  angle  is  equal  to  the  sine-of  its  supplement 
and  vice  versa,  and  AD  =  100 

AB  _  sin  30°  _       0.5 

100  ~ 
The  length  of  the  incline  is  then 


sin  10°      0.17365 


'-=287.93 


76 


SOME  PROBLEMS  IN  SURVEYING 


77 


The  depth  of  the  vertical  shaft  EC  is  now  easily  calculated  from 
the  right  triangle  ABC;  thus: 

BC  =  AB  X  sin  A  =  287.93  X  sin  20° 
=  287.93  X  0.34202  =  98.48  //. 

The  distance  DC,  from  the  mouth  of  the  drift  to  the  point  where 
the  upraise  should  be  started,  can  then  be  calculated  from  the 
triangle  ABC  by  subtracting  100  from  the  distance  AC  thus  found, 
as  follows: 

DC  =  AB  cos  20°  —  100  =  287.93  X  0.9397  ~~  I0°  =  I7°-57  ft- 
or,  directly  from  the  triangle  DBC,  thus: 


DC  = 


BC 


•  =  170.57  /*. 


tan  30      0.57735 

Problem  2. — (a)  If  the  course  of  a  main  entry  is  due  north,  what 
is  the  course  of  a  face  entry  turned  off  to  the  right,  at  an  angle  of 


FIG.    28.      PLAN    OF    MAIN,    FACE    AND    BUTT    ENTRIES,     AND    DIAGRAM    SHOWING 
THE   CORRESPONDING   COURSES. 

80° ?  (b)  What  is  the  course  of  a  butt  entry  turned  off  the  face 
entry  to  the  right,  at  an  angle  of  90°?  (c)  What  is  the  course  of  a 
room  turned  off  the  butt  entry,  at  an  angle  of  80°  to  the  right? 

(a)  In  Fig.  28  the  general  position  and  direction  of  the  main  en- 
tries, face  entries,  butt  entries  and  the  rooms  turned  off  the  butt 
entries  are  shown.     The  course  of  the  main  entries  being  due  north 
and  the  face  entries  being  turned  to  the  right  an  angle  of  80°,  the 
course  of  these  entries  will  lie  in  the  northeast  quadrant,  as  shown 
on  the  right  of  the  figure,  and  its  bearing  is  N  80°  E. 

(b)  The  butt  entries  being  turned  90°,  again,  to  the  right,  the 
azimuth  of  their  course  is  80  +  90  =  170°.     Since  this  azimuth  lies 


7  8  COAL  MINE  SURVEYING 

between  90°  and  180°,  the  course  of  the  butt  entries  lies  in  the 
southeast  quadrant.  All  bearings  in  the  southeast  and  southwest 
quadrants  being  estimated  from  the  south  end  of  the  meridian, 
the  angle  of  bearing,  in  this  case,  is  found  by  subtracting  the 
azimuth  from  180°.  Thus,  180  — 170  =  10°.  The  bearing  of  the 
butt  entries  is  then  S  10°  E. 

(c)  The  rooms  being  turned  80°  to  the  right  of  the  butt  entry, 
the  azimuth  of  the  rooms  is  170+80  =  250°.  Since  this  angle 
lies  between  180°  and  270°,  the  course  of  the  rooms  lies  in 
the  southwest  quadrant,  and  the  angle  of  bearing  measured  from 
the  south  end  of  the  meridian  is  250— 180  =  70°.  The  course 
of  the  rooms  is,  therefore,  S  70°  W. 

Problem  3. — An  approximate  method  of  measuring  across  a 
stream  by  use  of  a  transit  and  tape  but  without  reference  to  a  book 
of  tables. 

First  Method. — Where  it  is  possible  to  cross  over  the  stream, 
the  following  method  may  be  used,  which  will  give  approximate 
results:  Referring  to  Fig.  29,  set  up  the  instrument  at  A  and  sight 
to  a  point  B.  Then,  by  means  of  the  3,  4,  5  method,  often  called 
the  3,  7,  12  method,  set  off  the  right  angle  OBT.  To  do  this,  first 
line  in  the  point  0,  on  the  line  AB,  with  the  instrument,  making 
the  distance  OB  3  ft.  or  any  multiple  thereof,  and  place  a  surveying 
pin  at  O.  With  .B  as  a  center  and  a  radius  of  4  ft.,  describe  an  arc; 


0  3 


c 

FIG.    29.      METHOD    OF    MEASURING    DISTANCE    ACROSS    A    RIVER. 

and  with  O  as  a  center  and  a  radius  of  5  ft.,  describe  another  arc 
intersecting  the  first  at  T.  The  angle  OBT  will  then  be  a  right 
angle. 

By  the  3,  7,  12  method,  the  end  of  the  tape  is  fastened  at  0,  and 
the  tape  is  then  carried  around  the  triangle  OBT  and  back  to  0. 
The  distance  around  the  triangle  OBTO  is  12  ft.  Now,  holding  the 
end  and  the  i2-ft.  mark  at  0,  with  the  3-ft.  mark  at  B,  pull  out  the 
tape  with  a  pin  at  the  y-ft.  mark,  which  will  establish  the  point  T, 
making  the  angle  OBT  a  right  angle. 

With  the  instrument  at  A,  turn  off  the  angle  BAG  equal  to  5°  43', 


SOME  PROBLEMS  IN  SURVEYING  79 

and  line  in  the  point  C  on  the  line  BC.  Now,  since  the  tangent 
of  5°  43 '  is  o.i,  the  distance  AB  will  be  ten  times  the  distance  BC. 
Thus,  if  the  distance  BC  equals  50  ft.,  AC  will  be  10  X  50  =  500  ft. 
Second  Method. — When  it  is  not  possible  to  cross  to  the  opposite 
side  of  the  stream,  the  following  method  can  be  used:  Referring 
to  Fig.  30,  establish  the  line  CD  more  or  less  parallel  to  the  stream, 


FIG.    30.      ANOTHER   METHOD    OF   MEASURING   ACROSS   A   STREAM. 

and  another  line  EF  parallel  to  the  first,  and  at  any  distance  AO 
from  it.  Now,  select  a  well-defined  point  or  object,  B,  on  the 
opposite  bank  and  line  in  the  point  E  on  the  line  CB,  and  likewise 
the  point  F  on  the  line  DB.  Measure  CD  and  EF  carefully.  The 
distance  AB  is  then  calculated  as  follows: 


If  the  several  distances  are  as  indicated  in  Fig.  30,  the  calculation 
is  as  follows: 


-  —r-  X  ioo  = 
200  —  160 


X  100  =  500  ft. 


Problem  4. — If  the  backsight  BA  (Fig.  31),  is  N.  40°  E.,  1230 
ft.,  and  the  foresight  BC  is  S.  60°  E.,  3042  ft;  what  is  the  length 
and  bearing  of  the  closing  side  CA  ? 

Starting  from  A,  the  bearings  and  length  of  the  courses,  together 
with  the  latitudes  and  departures,  are  as  follows: 

Bearing  Distance  Latitude  Departure 

S  40°  W  1230  942.188  790. 64  W 

S  60°  E  3042  1521.00  S  2634. 37  E 

2463 . 18  S  1843 . 73  E 

The  total  latitude  or  southing  is,  therefore,  2463.18  ft.,  and  the  total 


go 


COAL  MINE  SURVEYING 


net  easting  1843.73  ft.     In  order  to  close  this  survey,  the  line  CA 
must,  therefore,  have  a  northing  of  2463.18  ft.,  and  a  westing  of 

1843-73  ft. 
To  find  the  bearing  of  this  closing  course,  call  the  angle  of  the 

bearing  a  ;  then, 

0.7485 


and  a  =  36°  49'.  The  bear- 
ing of  the  closing  course  is, 
therefore,  N.  36°  49'  W. 

The  length  of  the  closing 
course  is  then  found  as 
follows: 


2463.18 
cos  3 6°  49' 


2463.18  = 
0.80056 
3076.8  //. 


FIG.   31 

COORDINATE   SYSTEM  OF  CALCULATING 
CLOSING  COURSE. 


Problem  5.  —  A  certain 
seam  cuts  a  vertical  fault 
and  the  upthrow  is  found  to 

PLATE  OF  SURVEY,    SHOWING  THE   be  go  ft.         The  geam  beVOnd 
, 

the  fault  dips  at  the  rate  of 
4  in.  per  yard.     What  is  the 

length  of  a  drift  rising  r\  in.  per  yard,  that  will  cut  the  seam  beyond 
the  fault? 

The  combined  dip  of  the  seam  and  rise  of  the  drift  is  4  +  i-S  = 
5.5  in.  per  yard.  •  Assuming  horizontal  measurements  in  the  seam  and 
the   drift   alike,    the   horizontal 
length    of    the  drift,    measured 
from  the  fault  to  the  place  where 
it  cuts  the  seam,  would  be: 

60  X  12 

—  —  —  =  130.9  yd* 

5-5  FIG.  32 

Fig.  32  shows  the  relative  posi- 
tion of  the  fault,  the  seam  on  each  side  of  the  fault,  and  the  stone 
drift  driven  to  the  rise,  across  the  strata,  to  connect  the  seam  at 
the  fault  with  the  seam  beyond. 

Problem  6.  —  A  method  of  turning  off  an  entry  at  right  angles 
without  the  use  of  a  compass. 


SHOWING  A  VERTICAL  UPTHROW 
OF  60  FT. 


SOME  PROBLEMS  IN  SURVEYING 


Si 


Suspend  a  bob  from  each  of  the  two  respective  points  or  stations 
A  and  B  (Fig.  33)  of  the  entry  survey. 

Stretch  a  string  carefully  in  line  with  these  bobs  or  points.  By 
means  of  this  string,  line  in  the  points  0  and  P  opposite  the  centers 
of  the  respective  cross  entries.  Also  line  in  the  points  a  and  b, 
respectively,  at  any  convenient  equal  distances,  on  each  side  of  0, 
Then  from  these  points  a  and  b  as  centers  and  with  any  fixed  radius 
ac,  greater  than  aO,  describe  in  turn  the  intersecting  arcs  which 
determine  the  point  c  and  the  line  OX  at  right  angles  to  AB.  These 


yj j ^  Main  Entry 

0          b  Survey  Line     P 


—A 


COM.  A6E 

FIG.   33.      SHOWING   TWO    METHODS    OF    SETTING    OFF  A    RIGHT    ANGLE    WITHOUT 
USING    A   COMPASS. 

lines  can  often  be  laid  off  with  chalk  on  the  roof,  but  the  work 
requires  care. 

In  a  similar  manner  the  triangle  Pmn  may  be  laid  out,  using  the 
numbers  3,  4,  5  or  any  multiple  of  these  more  convenient,  as  6,  8, 
10  ft.  This  gives  a  right  triangle,  because 

62  +  82  =  io2 

and  makes  the  line  PY  at  right  angles  to  AB. 

It  is  important  to  remember  that  all  measurements  must  be 
made  in  the  horizontal  plane  for  any  angle  other  than  a  right  angle, 
which  can  be  laid  out  on  the  pitch. 


MINING  METHODS 

BY  ' 

ROBERT  BRUCE  BRIXSMADE,  B.  S.,  E.  M. 


COPYRIGHT,  1911,  BY  THE 
McGRAw-HiLL  BOOK  COMPANY,  INC. 


CONTENTS 

CHAPTER  I 
Explosives  and  Their  Use  in  Mining 

CHAPTER  II 

Principles  of  Blasting  Ground 

CHAPTER  III 
Compressed  Air  for  Mining 

CHAPTER  IV 

Principles  for  Controlling  Excavations 

CHAPTER  V 

Principles  of  Mine  Drainage 

CHAPTER  XVIII 


Principles  of  Mining  Seams      

(a)  Comparison  of  Longwall  and  Pillar  Systems . 

(b)  Comparison  of  Advancing  and  Retreating. 

(c)  Mining  by  Roof-pressure 


CHAPTER  XIX 

Advancing  Longwall  Systems  for  Seams 

Example  49. — Spring  Valley  Collieries,  111 

Example  50. — Montour  Iron  Mines,  Danville,  Pa.  . 
Example  51.— Bull's  Head  Colliery,  Eastern,  Pa.  . 
Example  52.— Vinton  Colliery,  Vintondale,  Pa.  .  . 
Example  53. — Drummond  Colliery,  Westville,  N.  S. 


CHAPTER  XX 

Pillar  Systems  for  Seams 

Example  54. — Advancing  System  Layouts 

Example  55. — Nelms'  Retreating  System 

Example  56.; — Nelms'  Advancing-retreating  System  .... 
Example  57.'- — Connellsville  District,  Western  Pennsylvania. 
Example  58. — Pittsburg  District,  Western  Pennsylvania  .  . 

CHAPTER  XXI 

Flushing  System  for  Filling  Seams  and  Recovering  Pillars  .  .  . 
Example  59. — Anthracite  District,  Eastern  Pennsylvania.  . 
Example  60. — Robinson  Mine,  Transvaal 


INDEX 


MINING  METHODS 


CHAPTER  I 
EXPLOSIVES  AND  THEIR  USE  IN  MINING 

An  explosion  may  be  defined  as  a  sudden  expansion  of  gas.  The 
substances  which  we  call  explosives  are  so  unstable  when  exposed  to  a 
suitable  flame  or  shock  that  they  suddenly  change  into  many  times 
their  original  volume  of  gas  with  the  evolution  of  heat.  If  the  change  to 
a  gas  takes  place  in  the  open,  there  is  a  flame  and  a  whiff  or  a  report.  It 
is  only,  however,  when  explosives  are  set  off  in  confined  spaces  like  drill- 
holes that  they  do  their  chief  work  in  mining.  Consequently  a  blast  or 
explosion  may  be  said  to  be  a  rapid  combustion  in  a  confined  space. 

Explosives  have  two  essential  constituents,  namely,  combustibles 
and  oxidizers.  They  may  be  broadly  divided  into  three  classes  accord- 
ing to  the  relation  which  the  combustibles  bear  to  the  oxidizers.  Class  I 
includes  the  mechanical  explosives,  or  those  in  which  the  ingredients 
constitute  a  mechanical  mixture;  class  II  includes  the  chemical  explo- 
sives or  those  in  which  the  ingredients  are  in  chemical  combination;  class 
III  includes  the  mechanico-chemical  explosives  which  are  formed  of  a 
mixture  of  class  II  and  an  absorber. 

METHODS  OF  FIRING  EXPLOSIVES 

Explosives  are  set  off  by  two  means — ignition  and  detonation. 
Because  through  ignition  the  combustion  is  transmitted  by  heat  alone, 
it  gives  a  slower  explosion  than  one  started  by  detonation  which  trans- 
mits the  reaction  by  the  rapidity  of  vibrant  motion.  By  their  nature 
class  I  is  adapted  to  ignition,  and  classes  II  and  III  to  detonation. 

Ignition  is  commonly  performed  by  squibs,  fuse  or  electric  igniters. 
A  squib  is  really  a  self -impelling  slow  match,  made  by  filling  one-half  of  a 
thin  roll  of  paper  with  black  powder  and  the  other  half  with  sulphur. 
For  their  use  in  blasting,  a  drill-hole  ab,  Fig.  1,  is  loaded  with  an  explo- 
sive be  and  before  filling  the  hole  with  the  tamping  cd,  a  needle  ac  is 
inserted  into  the  explosive  so  that  when  it  is  withdrawn,  a  hole  of  a 
larger  diameter  than  the  squib  is  left  through  the  tamping  from  a  to  c. 

1 


MINING    WITHOUT    TIMIJKIl 


Fio.   1. — Drill-hole  section. 


The  squib  is  then  inserted  in  this  hole  with  the  sulphur  end  out,  and 
when  lit  the  slow-burning  sulphur  allows  time  for  the  miner  to  escape 
before  the  powder  of  the  squib  takes  fire  and  its  reaction  forces  the  squib 
along  the  holes  to  ignite  the  powder  at  c. 

A  fuse  is  merely  a  thread  of  black  powder  wrapped  with  one  or  more 
thicknesses  of  tape.  In  loading  the  hole,  Fig.  1,  the  fuse  would  be 
inserted  in  place  of  the  needle  ac.  A  fuse 
burns  commonly  at  the  rate  of  2  ft.  a 
minute.  Therefore  a  sufficient  length 
should  be  used  in  the  hole  to  allow  the 
miner  to  retire  in  safety,  after  splitting  and 
lighting  the  outer  end,  before  the  flame 
reaches  the  explosive  at  c. 

The  electric  igniter  consists  of  a  shell  a, 
Fig.  2,  enclosing  a  charge  of  fulminate 
mixture  in  b  and  of  sulphur  cement  in  e. 
The  copper  wires  c  pass  through  /  and 
enter  6  where  they  are  connected  by  a 
platinum  bridge  at  d.  For  ignition,  the 
shell  a  is  made  of  pasteboard  and  the 
igniter  is  placed  within  the  explosive  while 
the  wires  extend  outside  the  hole  to  a 
blasting  machine.  The  last  is  simply  a  small  armature  revolving 
between  its  poles  and  sending  a  current  through  the  igniters  in  the 
circuit  when  its  handle  is  shoved  down.  All  the  common  electric 
igniters  on  one  circuit  are  exploded  simultaneously,  but  a  recent  inven- 
tion is  a  delay-action  igniter  which  permits  electric  firing  in  sequence. 

Detonation  is  performed  by  fuse  and  cap  or  by  electric  caps.  A 
blasting  cap  is  simply  a  cylindrical  copper  cup  with  a  small  charge  of 
fulminate  mixture  in  its  bottom,  the  fuse  being  inserted  into  the  cup  and 
fastened  to  it  by  crimping  pincers.  The  cap  is  then  inserted  into  one 
cartridge  of  the  explosive  and  its  attached  fuse  tied  firmly  to  it  by  a 
string,  in  order  to  make  a  primer 
which  is  placed  near  or  on  the  top  of 
the  explosive.  The  loaded  hole  will 
then  resemble  Fig.  1,  the  explosive 
being  in  be,  the  cap  and  primer  at  c, 
and  the  fuse  along  ca.  Lighting  the 
fuse  is  the  same  as  for  ignition,  only  the  fuse  now  fires  the  cap  whose 
explosion  detonates  the  explosive. 

The  electric  cap  resembles  the  electric  igniter,  Fig.  2,  but  has  a  copper 
instead  of  a  pasteboard  case  a  and  the  quantity  of  charge  of  fulminate 
mixture  at  b  is  increased  as  the  sensitiveness  of  the  explosive  diminishes. 
The  electric  cap  is  inserted  in  and  fastened  to  a  primer-cartridge  like 


Fio.  2. — Electric  exploder. 


EXPLOSIVES    AND    THEIR    USE    IN    MINING 

fuse  and  cap,  the  electric  cap  being  fired  by  a  blasting  battery  in  the 
same  way  as  the  electric  igniter. 

LOADING  AND  TAMPING 

A  mechanical  explosive  like  black  powder  usually  comes  in  bulk. 
For  loading  it  is  poured  into  a  cartridge  (the  size  of  the  hole)  which  is 
made  by  rolling  a  piece  of  paper  around  a  pick  handle.  For  damp 
holes  the  cartridge  must  be  oiled  or  soaped  on  the  outside.  This  paper 
cartridge  is  pressed  down  into  the  hole  by  a  soft  iron  tamping  bar  whose 
tip  should  be  an  expanding  copper  cone  grooved  on  the  edge  for  the 
purpose  of  allowing  the  copper  loading  needle  or  fuse  to  pass.  Tamping 
bars  with  iron  tips  or  iron  needles  are  highly  dangerous  in  formations 
containing  pyrite  or  other  hard  minerals,  on  which  the  iron  might  strike 
a  spark,  and  their  use  is  therefore  prohibited  by  law  in  many  places. 

A  mining  explosive  of  class  II  or  III  is  handled  in  paper  cartridges 
which  can  be  ordered  of  a  diameter  to  fit  the  hole.  Before  loading  they 
are  slit  around  lengthwise  to  permit  of  the  explosive  taking  the  shape  of 
the  hole  when  it  is  pressed  down  by  a  tamping  bar  which  should  be  of 
wood  for  these  explosives,  instead  of  copper-tipped  iron,  on  account  of 
their  being  more  sensitive  to  any  shock  than  black  powder. 

In  coal  mines,  coal  dust  is  commonly  used  for  tamping  black  powder, 
but  this  is  a  very  unsafe  practice  in  dangerous  mines,  for  a  windy  or 
blown-out  shot  will  have  its  normal  flame  increased,  both  in  length  and 
duration,  by  the  ignition  of  the  tamping.  The  best  materials  for  tamp- 
ing are  a  fine  plastic  clay  or  loam  and  ground  brick  or  shale,  and  al- 
though sand  is  too  porous  to  do  well  for  black  powder,  it  answers  for 
higher  explosives  but  must  be  confined  in  paper  cartridges  for  use  in 
uppers. 

Water  is  used  as  tamping  for  nitro-glycerine  and  high  explosives  in 
wet  down-holes,  but  it  is  little  better  than  nothing.  The  fact  that  higher 
explosives  will  break  rock  without  any  tamping  has  caused  many  miners 
to  abandon  tamping  them  altogether  on  account  of  the  ease  of  recapping 
untamped  charges  in  case  of  a  misfire.  Mechanical  explosives  must  be 
tightly  tamped,  nearly  to  the  collar  of  the  hole,  or  they  will  blow  out 
instead  of  breaking  the  rock,  and  although  the  tamping  may  be  shortened 
with  detonating  explosives,  as  they  become  quicker  and  stronger,  a  short 
length  of  tamping  adds  to  the  efficiency  of  the  highest  explosives. 

Where  only  quick-acting  explosives  of  classes  II  or  III  are  at  hand 
and  it  is  desired  to  blast  with  the  slow  action  of  class  I,  the  object  can 
be  partially  obtained  by  special  methods  of  loading.  These  methods 
provide  an  air  cushion  between  the  explosive  and  the  rock  and  tamping 
by  either  having  the  stick  of  explosive  of  considerably  smaller  diameter 
than  the  drill  hole  or  by  having  a  very  porous  cellular  tamping  to  sepa- 
rate the  tight  tamping  from  the  explosive. 


4  MINING    WITHOUT    TIMBER 

Before  examining  the  various  mine  explosives  in  detail,  let  us  consider 
an  illustration  of  the  method  of  calculating,1  from  the  chemical  equation 
of  an  explosive,  its  calorific  power,  its  temperature,  and  the  number  of 
expansions  and  its  consequent  exploding  pressure.  Let  us  assume  the 
simplest  case  of  a  mechanical  mixture  of  hydrogen  and  oxygen  at  a 
temperature  of  0°  C.  and  at  sea-level  pressure  of  760  mm.  of  mercury. 
Then  the  chemical  equation  for  complete  combustion  is 

2H2  +  02  =  2H20.  (1) 

the  molecular  weights  being  4  +  32  =  36.  (2) 

If  t  =  thermometer  temperature  in  degrees  centigrade  of  the  explosion; 
T  =  absolute  temperature  in  degrees  centigrade  of  the  explosion; 
S  =  sign  for  summation; 

^  etc.  =  weights  in  grams  of  various  combustibles  of  the  explosive; 
2,  etc  =  calorific  power  in  calories  of  various  products  of  combustion 
of  the  explosive; 
wwlw3,  etc.  =  weights  in  grams  of  various  products  of  combustion  of  the 

explosive; 

sSjSj,  etc.  =  specific  heat  in  calories  of  various  products  of  combustion  of 
the  explosive; 

V  =  volume  of  explosive  originally; 

VI  —  volume  of  explosive  due  to  chemical  reaction  alone; 

V2  =  volume  of  explosive  due  to  chemical  reaction  and  resulting  tempera- 

ture, t; 

P  =  pressure  of  explosive  originally; 
P2  =  pressure  of  explosive  finally; 
then  we  have  from  thermo-chemistry, 

_  WC  +  TFA  +  TF2C2,  etc.  =  SWC 


For  the  given  problem  we  have  from  equation  (2) 

W  =  4:  grams  of  H  gas; 

w  =  36  grams  of  H20  vapor. 
From  thermo-chemistry  we  have, 

C  =  28,780  cal.  for  H; 

s  =  0.4805  cal.  for  H2O  vapor; 
substitute  in  (3)  and 

-      4X28'78°-=6660°C. 


36X0.4805 

Then,  from  Avogardro's  law,  that  the  molecules  of  equal  volumes  of  all 
gases  under  like  conditions  occupy  the  same  volume,  we  have  from  (1), 

2  vols.  H  +  l  vol.  O  =  2  vols.  H2O, 
or 

Vl  =  2/3V.  (4) 

i  See  "Metallurgical  Calculations,"  by  J.  W,  Hichards. 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  5 

From  Charles'  law,  the  volumes  of  gases  vary  directly  as  their  abso- 
lute temperature  we  have  thus 

!>=        T 
Vl     0  +  273 

or 

_  6660  7t< 
2~~273~' 
substitute  from  (4)  and  we  have 

»- 


From  Boyle's  law,  if  the  gas  of  volume  72  is  prevented  from  expand- 
ing beyond  volume  7,  we  have  for  the  final  pressure  P2  in  the  explosive 
chamber  P, 


or 

fa  =  f'*  (6) 

Substitute  in  (6)  from  (5)  and,  as    P  =  1  atmosphere  =  14.7  Ibs.  per 
sq.  in.,  we  have 

16  27  P 
PJJ=  —  -  -  -=16.2   atmospheres. 

or  238  Ibs.  per  sq.  in 

From  physics,  T  =  t  +  273, 
hence 

t  =  T—  273  =  6660—273  =  6387°  C. 

In  practice,  this  theoretical  pressure  and  temperature,  resulting  from 
the  explosion,  would  have  to  be  multiplied  by  a  fractional  factor  of 
efficiency  to  allow  for  imperfect  combustion  and  loss  of  heat  through 
radiation  and  leakage.  In  large  charges,  these  losses  are  proportionally 
less  than  in  the  case  of  small  charges.  This  fact,  coupled  with  the  greater 
likelihood  of  their  meeting  weak  places  in  the  blast's  burden,  accounts 
for  the  higher  efficiency  of  the  former.  These  theoretical  calculations 
are  especially  useful  in  comparing  the  relative  strength  of  different 
explosives  of  the  same  type.  In  France,  they  are  used  extensively  in  the 
inspection  of  permissible  explosives  to  determine  if  their  final  tempera- 
ture is  sufficiently  low  for  use  in  dangerous  coal  mines. 

The  practical  usefulness  of  explosives  depends  upon  (1)  their  cost  of 
manufacture;  (2)  their  safety  and  convenience  as  regards  transportation 


6  MINING    WITHOUT   TIMBER 

and  storage;  (3)  method  necessary  for  their  loading  and  exploding; 
(4)  their  exploding  pressure;  (5)  the  rapidity  with  which  they  explode; 
(6)  the  length  and  temperature  of  the  flame.  These  six  factors  will  now 
be  discussed  seriatim.  Factor  (1),  or  the  cost,  is  often  the  most  impor- 
tant factor  in  commercial  operations  like  mining,  although  for  purposes 
of  war  it  is  often  little  considered.  Factor  (2)  or  safety,  affects  the 
desirability  for  all  purposes,  the  more  sensitive  the  explosive,  the  higher 
the  freight  rate  by  rail  or  boat,  and  if  sensitive  beyond  a  certain  point, 
it  cannot  be  shipped  thus  at  all.  Those  explosives  which,  like  dynamite, 
freeze  at  ordinary  winter  temperatures  are  at  a  disadvantage  as  are  also 
those  which,  like  black  powder,  are  handled  loose  and  can  be  easily 
ignited  by  a  spark  struck  by  a  hob-nailed  shoe  on  a  floor  spike.  Some 
explosives,  like  imperfectly  washed  guncotton,  are  liable  to  explode  by 
spontaneously  generated  heat,  while  others  become  dangerously  sensi- 
tive if  exposed  to  the  sun  during  shipment.  The  desirability  of  explo- 
sives belonging  to  either  of  these  last  two  mentioned  classes  is  plainly 
discounted  because  of  these  attributes.  The  next  factor  (3)  or  loading 
and  exploding,  is  important  in  connection  with  conditions  such  as  prevail 
in  dangerous  coal  mines  (where  an  open  light  is  prohibited),  in  subaque- 
ous blasting  (where  both  explosive  and  exploder  must  be  unaffected  by 
water),  or  where  misfires  could  not  be  corrected.  Factor  (4),  or  the 
pressure,  is  what  determined  the  real  effective  breaking  force  of  the 
explosion,  but  it  is  modified  in  practice  by  (5),  or  the  rapidity  of  the 
explosion.  Slow  and  fast  explosives  are  comparable  to  presses  and 
hammers  for  forging  steel.  The  former  exerts  its  pressure  gradually 
until  the  strain  exceeds  the  tensile  strength  of  the  material  and  the  rock 
gives  way  along  a  surface  of  fracture.  The  latter  gives  a  sharp  quick 
blow  which  will  shatter  the  surface  of  rock  exposed  to  the  explosive  before 
any  fracturing  action  is  exerted  on  the  blast's  burden  of  rock. 

The  slow  explosive  will  detach  the  rock  in  large  masses  while  the 
fast  type  may  crush  it  to  bits.  Black  powder  is  an  example  of  the  first 
and  nitro-glycerine  of  the  second.  Explosives  with  all  graduations  of 
rapidity  between  these  extremes  are  on  the  market.  The  fastest  explo- 
sives are  applicable  where  the  rock  is  very  hard  to  drill  as,  for  example, 
in  the  case  of  certain  Lake  Superior  hematites,  or  where  a  tremendous 
force  must  be  exerted  from  confined  spaces  as  in  breaking  the  cut  for 
development  passages;  also  where  a  shattering  rather  than  a  fracturing 
action  is  needed,  as  in  chambering  the  bottom  of  drill  holes  or  in  shooting 
oil  wells.  The  slowest  explosives  are  used  in  quarrying,  for  the  purpose 
of  detaching  monoliths,  or  in  consolidated  or  soft  rock  which  can  be 
fractured  by  a  slow,  pressing  movement  but  only  dented  by  a  quick 
hammer  blow. 

Factor  (6),  or  the  flame  and  temperature,  is  an  important  considera- 
tion for  blasting  in  gassy  or  dusty  coal  mines.  The  so-called  "  permis- 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  7 

sibles"  are  explosives  made  to  fall  below  a  minimum  legal  requirement 
as  regards  length  and  temperature  of  flame.  When  one  considers  that  a 
permissible  like  carbonite  gives,  in  practice,  a  flame  height  of  15.8  in. 
and  a  flame  duration  of  0.0003  seconds,  as  compared  with  50.2  in.  and 
0.1500  seconds  respectively,  for  black  powder,  we  can  see  how  much 
safer  the  permissible  is  to  use. 

We  will  now  consider  the  properties  of  the  three  classes  of  explosives: 

CLASS  I,  OR  MECHANICAL  EXPLOSIVES 

The  common  representatives  of  this  class  are  black  powder  and 
mechanical  permissible  explosives.  Black  powder  was  discovered  before 
600  A.  D.  by  the  Chinese,  and  by  Roger  Bacon  in  1270,  but  it  was  not 
used  for  mining  until  Martin  Weigel  introduced  it  at  Freiberg  in  1613. 
It  can  be  made  from  a  single  combustible,  charcoal,  mixed  with  an 
alkaline-nitrate  oxidizer,  but  in  order  to  lower  its  ignition  temperature 
for  blasting  to  about  275°  C.,  part  of  the  charcoal  is  replaced  by  sulphur. 
For  the  cheaper  blasting  powders,  the  oxidizer  is  sodium  nitrate  which, 
being  easily  affected  by  dampness,  is  replaced  in  the  higher  grade  powders 
by  potassium  nitrate.  The  ingredients  are  first  ground  then  mixed 
thoroughly  while  moist  and  finally  pressed  in  cakes,  dried,  broken  and 
sized.  Assuming  the  equation  for  the  complete  combustion  of  black 
powder  to  be. 

3C  +  S  +  2KN08=3C02  +  N  +  K2S.  (7) 

We  have  by  calculation  for  its  percentage  composition, 

carbon  =  13. 4 

sulphur  =11.8 

sodium  nitrate  =  74. 8 

100^0 

and  for  the  percentage  composition  by  volume  of  its  resulting  gas, 

C02  =  75 

N     =25 

~100~ 

The  theoretical  exploding  temperature  is  4560°  C.  and  the  pressure  is 
5820  atmospheres.  In  practice  the  composition  is  varied  according  to 
the  experience  of  each  maker.  As  the  combustion  is  imperfect,  poison- 
ous and  combustible  gases  like  carbon  monoxide,  hydrogen  sulphide 
and  hydrogen  and  unpleasant  vapors,  like  the  sulphide,  sulphate,  hypo- 
sulphite, nitrate  and  carbonate  of  potassium,  are  given  off  by  the  explo- 
sion and  sometimes  render  breathing  or  the  carrying  of  open  lights  in 
the  fumes  a  dangerous  procedure.  In  fact,  Bunsen's  experiments  proved 


8  MINING    WITHOUT   TIMBER 

that  only  one-third  of  the  ignited  gunpowder  really  followed  the  reaction 
of  equation  (7). 

Black  powder  is  sold  in  grains  which  vary  in  size  from  the  fine  sporting 
gunpowder  to  the  2-in.  balls  of  artillery  powder.  For  blasting,  the  grains 
vary  in  diameter  from  one-eighth  to  one-half  of  an  inch,  and  the  rapidity 
of  the  explosion  decreases  with  an  increased  diameter  of  grain.  The 
grains  should  be  of  uniform  size,  quite  dry  and  thoroughly  tamped  in 
the  hole  in  order  to  get  good  results.  The  specific  gravity  of  lightly 
shaken  black  powder  is  about  the  same  as  water.  Its  cheapness,  non- 
freezing,  comparative  safety  for  shipping  and  handling,  easy  explosion 
by  ignition  and  slow  action  are  the  favorable  qualities  of  black  powder 
which  cause  its  wide  use.  For  coal  mines  free  from  dangerous  gases  and 
dust,  it  is  a  better  explosive  than  detonating  permissibles  whose  quicker 
action  breaks  up  the  coal  and  injures  the  roof  more.  Black  powder  is 
rendered  inefficient  for  many  other  purposes,  however,  because  of  its 
necessitating  much  tamping,  its  low  power,  the  readiness  with  which  it 
is  spoiled  by  moisture  and  its  long  flame. 

Of  the  mechanical  permissibles  bobbinite  has  been  extensively  used 
in  England.  Its  percentage  composition  is, 

Potassium  nitrate  =  65.0 

Charcoal  =  20 . 0 

Sulphur  =  2.0 

Paraffin  wax=   2.5 

Starch  =   8.0 

Water  =   2.5 


100.0 

It  is  thus  chemically  very  close  to  black  powder  excepting  that  it 
contains  more  charcoal  and  less  sulphur  and  makes  up  that  discrepancy 
by  the  addition  of  wax,  starch  and  water.  The  lack  of  sulphur  raises  its 
ignition  temperature  while  the  wax  forms  a  waterproof  coating  for  the 
grains  of  powder.  The  starch  and  water  absorb  heat,  shorten  the  flame 
and  decrease  the  exploding  temperature  to  under  1500°  C.  It  is  handled 
in  compressed  cartridges  with  wax  coverings.  It  has  a  central  hole  to 
admit  the  fuse,  for  ignition  by  squib  is  not  allowed  in  dangerous  coal 
mines. 

CLASS  II,  OR  CHEMICAL  EXPLOSIVES 

The  five  common  explosives  of  this  class  are  guncotton,  nitro-glycerine, 
nitro-gelatin,  fulminates  and  picrates.  They  all  contain  nitryl  (NO2) 
and  their  detonation  is  made  possible  by  the  unstable  quality  of  nitryl 
compounds. 

Guncotton, — This  was  discovered  by  Schonbein  in  1846,  but  it  was 
little  used  until  it  was  found  that  its  dangerous  instability  was  not 


EXPLOSIVES    AND    THEIR    USE    IN"    MINING  9 

inherent  but  due  solely  to  the  surplus  acid  left  in  its  tissue  by  imperfect 
washing  methods  during  its  manufacture.  The  equation  for  making 
it  is, 

C6H1005  +  3HN03  =  C6H706  (NO2)  3 + 3H20.  (8) 

cotton  +  nitric  acid  =  guncotton  +  water. 

The  ingredients  are  allowed  to  stand  in  a  cold  place  for  some  time 
before  the  washing  out  of  the  free  acid  is  begun. 
The  reaction  on  exploding  is, 

2C6H705  (N02)  3 = 3C02  +  9C02 + 3  N2  +  7H20.  (9) 

Equation  (9)  shows  that  the  explosion  gives  no  solid  product  like  the 
K2S  of  equation  (7)  and  that  the  percentage  composition  by  volume  of 
the  resulting  gas  is, 

CO2  =  13.7 

CO  =  40.8 

N  =  13.7 

H20=31.8 

100.0  f 

By  the  method  of  calculation  already  explained,  it  is  found  that 
guncotton  theoretically  has  an  exploding  temperature  of  5340°  C.  and  a 
pressure  of  20,344  atmospheres. 

The  combustible  qualities  of  the  large  percentage  of  carbon  monoxide 
resulting  from  its  explosion  render  guncotton  unfit  for  use  in  coal  mines, 
and  its  poisonous  qualities  make  it  unsuitable  for  any  underground  use. 
For  surface  work,  it  is  very  powerful,  smokeless,  does  not  freeze  and  is  not 
volatilized  or  decomposed  by  atmospheric  temperature.  It  ignites 
between  270  and  400°  F.  and  if  unconfined  will  then  burn  quietly.  When 
dry,  it  is  sensitive  to  percussion  and  friction,  but  under  water  it  is  insen- 
sible to  ordinary  shocks.  Immersed,  it  absorbs  from  10  to  15  per  cent, 
of  water,  but  even  then  it  can  be  exploded  without  drying  by  the  use  of 
an  extraordinarily  strong  detonator.  Its  chief  disadvantage  above 
ground  is  its  high  cost  and  the  fact  that  it  comes  in  hard  compressed 
cartridges  (specific  gravity  about  1.2)  which  fit  drill  holes  only  imper- 
fectly and  therefore  lose  in  efficiency.  For  any  destructive  work  without 
the  use  of  drill  holes,  like  demolishing  walls,  dams  and  the  like,  the  sharp, 
sledge-hammer  blow  of  its  explosion  renders  it  very  efficacious. 

Nitro-glycerine  or  "Oil." — This  was  discovered  by  Sabrero  in  1847, 
but  did  not  become  commercially  valuable  until  1863  under  the  direction 
of  Alfred  Nobel.  The  equation  for  its  making  is, 

C3H803+3HN03  =  C3H503(N02)3+3H20.  (10) 

glycerine  +  nitric  acid  =  nitro-glycerine  +  water. 

Strong  sulphuric  acid  is  an  ingredient  of  the  mixture,  but  it  does  not 
take  part  in  the  reaction,  which  must  take  place  at  a  moderate  tempera- 


JO  MINING    WITHOUT    TIMBER 

ture  to  be  safe.  The  resulting  "oil"  is  much  easier  to  wash  than  gun- 
cotton  and  consequently  is  cheaper.  It  is  a  yellow,  sweetish  liquid 
poisonous  both  to  the  blood  and  the  stomach.  Its  specific  gravity  is  1.6. 
Its  freezing-point  is  about  45°  F.  and  to  insure  against  freezing  the  tem- 
perature must  be  above  52°  F.  When  frozen,  it  is  insensible  to  ordinary 
shocks,  as  is  also  the  case  when  it  is  dissolved  in  alcohol  or  ether.  It  is, 
therefore,  commonly  shipped  either  in  tin  cans,  packed  in  ice,  or  in 
solution  hi  wood  alcohol.  It  can  be  precipitated  from  the  latter  before 
use  by  an  excess  of  water. 

Nitro-glycerine  does  not  evolve  nitrous  fumes  until  230°  F.  As  it 
begins  to  vaporize  at  about  100°  F.,  it  is  important  in  thawing  it  not  to 
exceed  this  temperature.  Thawing,  therefore,  is  only  safely  done  by 
heating  the  explosive  over  a  water  bath  at  less  than  90°  F.,  or  by  leaving 
it  in  a  room  of  the  same  temperature  for  some  time.  The  explosive 
ignites  at  only  356°  F.  and  if  then  pure  and  free  from  all  pressure,  jar  or 
vibration,  it  will  burn  quietly.  These  safe-igniting  conditions,  however, 
are  difficult  to  obtain,  for  a  small  depth  of  liquid  causes  sufficient  pressure 
to  explode  it  when  ignited.  Thus  a  film  of  it,  heated  on  a  tin  plate, 
burned  without  an  explosion  only  if  under  one-fourth  inch  thick.  The 
exploding  temperature  is  380°  F.  This  24°  margin  above  the  igniting 
temperature  accounts  for  the  numerous  cases  of  conflagration  without 
explosion.  The  reaction  of  the  explosion  is, 

4C3H603(N02)  3  =  12C02  +  02  + 3N2  +  10H2O.  (11) 

From  equation  (11)  the  explosive  product  is  gaseous  and  its  percen- 
tage composition  by  volume  is 

CO2=  46.0 

0=     3.8 

N     =   11.8 

H20=  38.4 

100.0 

By  the  previous  calculating  method,  it  is  found  that  theoretically 
the  exploding  temperature  is  6730°  C.  and  the  pressure  is  29,107  atmos- 
pheres. From  the  fact  that  its  explosive  product  contains  no  carbon 
monoxide,  "oil"  can  be  used  underground,  but  only  when  mixed  with 
an  absorber.  Alone,  it  is  too  sensitive  to  be  safe,  while  being  liquid,  if 
unconfined,  it  would  leak  from  holes  in  porous  rock,  and  if  confined  in 
canisters  it  will  not  fill  the  drill  hole.  With  its  great  speed  and  strength 
it  also  tends  to  shatter  locally  any  enclosing  rock,  except  the  toughest, 
rather  than  detach  it.  These  characteristics  render  it  inefficient  for 
most  mining  work. 

For  shooting  oil  wells,  however,  its  shattering  quality  renders  it 
peculiarly  suitable.  For  this  purpose,  a  cylindrical  canister  of  a  diam- 
eter to  fit  the  well  and  containing  from  100  to  200  Ibs.  of  nitro-glycerine, 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  11 

is  carried  to  the  well  swung  from  the  body  of  a  spring  buggy.  After 
filling  the  well  with  water,  the  canister  is  topped  with  a  cap  and  lowered 
to  the  proper  depths  by  a  rope,  along  which  a  weight,  called  a  "  go-devil," 
is  dropped  onto  the  cap  to  cause  the  explosion. 

Nitro-gelatin.  —  This  was  discovered  by  Nobel  in  1875  and  is  a  yellow- 
ish jelly  of  considerable  toughness,  but  easily  cut  with  a  knife.  It  is 
made  by  dissolving  guncotton  in  nitro-glycerine.  Authorities  differ  in 
the  proportion  of  guncotton,  some  recommending  only  7  per  cent.  To 
balance  all  the  free  oxygen  of  the  nitro-glycerine  by  the  excess  carbon 
of  the  guncotton  alone,  takes  87.3  per  cent,  of  the  former  to  12.7  per  cent. 
of  the  latter  and  gives  the  following  equation: 

9C3H503(NO2)3  +  C6H702(N02)3=33C02  +  15N2  +  26H20.       (12) 

From  equation  (12)  the  percentage  composition  of  the  solely  gaseous 
product  is, 

C02  =  44.6 
N=  20.2 
H20=35.2 

By  the  theoretical  calculation,  the  exploding  temperature  is  7080°  C. 
and  the  pressure  is  27,100  atmospheres.  The  last  figure  shows  nitro- 
gelatin  to  be  only  7  per  cent,  weaker  by  weight  than  nitro-glycerine, 
while  its  somewhat  higher  cost  is  due  to  its  guncotton  ingredient.  When 
used  alone  for  military  purposes,  about  4  per  cent,  of  camphor  is  dissolved 
in  the  nitro-glycerine  along  with  the  guncotton  to  make  a  product  called 
military  gelatin.  The  last  explosive  is  so  insensitive  that  it  can  be 
punctured  without  effect  by  a  rifle  bullet.  The  common  nitro-gelatin 
is  much  less  sensitive  than  No.  1  dynamite,  to  shock  or  friction,  and 
unaffected  by  a  short  immersion  in  water  at  158°  F.  and  by  an  8-day 
immersion  at  113°  F. 

It  will  not  exude  nitro-glycerine  under  a  high  pressure  or  any  atmos- 
pheric temperature.  Its  specific  gravity  is  1.6  and  it  can  be  set  off  only 
by  a  strong  detonation.  It  ignites  at  399°  F.  and  will  then  only  burn 
when  unconfined.  When  it  freezes,  which  is  between  35  and  40°  F.,  it 
becomes  more  sensitive  than  normally  owing  probably  to  the  partial 
freeing  of  the  nitro-glycerine  ingredient. 

Nitro-gelatin  is  now  used  for  mining  wherever  the  highest  power 
explosive  is  needed  and  is  especially  adapted  to  wet  or  subaqueous 
blasting,  either  alone  or  as  "gelatin"  dynamite. 

Fulminates.  —  Mercuric  fulminate  is  the  common  commercial  salt. 
It  is  made  as  follows  from  mercuric  nitrate  and  alcohol: 


(13) 
The  explosive  reaction  is 

Hg(CNO)2  =  Hg+2CO+2N.  (14) 


12  MINTING    WITHOUT    TIMBER 

Equation  (14)  shows  that  mercuric  fulminate  is  a  poor  explosive 
because  it  produces  the  poisonous  fumes  of  Hg  and  CO  as  well  as 
unburned  carbon.  If  a  little  damp,  it  explodes  very  feebly  and  if  quite 
wet,  not  at  all.  However,  its  non-freezing  quality,  its  quick  hammer- 
like  vibrant  explosion  and  its  uniform  sensitiveness  to  ignition  or  shock 
cause  its  use  as  the  chief  ingredient  of  percussion-cap  mixtures  for  deton- 
ating other  explosives.  Its  exploding  temperature  is  305°  F. 

Picrates. — These  salts  are  founded  on  picric  acid,  which  is  made  by 
mixing  carbolic  and  nitric  acid  according  to  the  equation, 

C6H60+3HN03  =  C6H3(N02)30+3H20.  (15) 

Its  explosive  reaction  is 

CaH3(N02)30  =  H2O+H  +  6CO  +  3N.  (16) 

Picric  acid  comes  in  yellow  crystals  which  are  soluble  in  hot  water  or  al- 
cohol, and  melt  at  230°  F.  It  is  used  very  largely  in  dyeing.  It  is  expens- 
ive to  make  and  difficult  to  explode.  Equation  (16)  indicates  that  it 
produces  much  of  the  poisonous  carbon  monoxide  which  shows  incom- 
plete combustion  and  consequently  a  decreased  power.  Picrates  are 
the  basis  of  the  military  explosive  lyddite,  but  the  recent  commercial 
failure  of  the  excellent  mining  picrate  "  joveite"  may  discourage  future 
attempts  to  adapt  them  to  blasting. 

CLASS  III,  MECHANICO-CHEMICAL  EXPLOSIVES 

This  class  will  be  considered  under  five  groups:  (1)  guncotton;  (2) 
nitro-glycerine;  (3)  nitro-gelatin;  (4)  fulminate;  (5)  nitro-benzol.  Deto- 
nating permissibles  for  coal  mining  fall  mainly  under  groups  (2)  and  (5) 
and  will  be  considered  last. 

Guncotton  Group. — The  evaporating  of  guncotton,  after  it  has 
been  dissolved  in  a  suitable  solvent  such  as  alcohol  or  acetone,  produces 
a  hard,  horny  material  which  is  the  basis  of  most  modern  smokeless  gun- 
powder. Its  chief  blasting  powder,  however,  is  tonite  which  is  formed 
by  adding  enough  barium  nitrate  to  guncotton  to  just  completely  oxidize 
the  gases  caused  by  the  explosion  as  follows: 

10CGH7O5(NO2)3  +  9Ba(NO3)2  =  60CO2  +  24N2+35H2O+9BaO.   (17) 

The  percentage  composition,  by  volume,  of  the  gaseous  product  of 
equation  (17)  is, 

C02=  50.4 

N=   20.2 

H2O=   29.4 

Tool) 

By  calculation,  the  exploding  temperature  is  5,730°  C.  and  the  pres- 
sure is  13,840  atmospheres,  which  are  fifteen-fourteenths  and  two-thirds, 


EXPLOSIVES   AND   THEIR    USE    IN    MINING  13 

respectively,  of  the  corresponding  figures  for  guncotton.  As  an  offset 
to  lessened  powder  tonite  is  plastic,  cheaper  than  guncotton  and  50  per 
cent,  denser.  Its  harmless  fumes  adapt  it  to  underground  use  and,  like 
dynamite,  it  is  packed  in  paper  cartridges.  It  has  been  extensively  used 
in  England,  where  it  is  shipped  under  the  same  safety  regulations  as 
black  powder.  It  is  hard  to  ignite  and  when  alight,  it  normally  burns 
slowly  without  explosion.  Tonite,  like  guncotton,  is  non-freezable  and 
is  detonated  only  by  a  strong  cap.  Potassium  nitrate  has  been  used, 
instead  of  barium  nitrate,  as  the  oxidizer,  in  another  guncotton  mixture 
of  similar  properties  which  is  called  potentite. 

Nitro-glycerine  Group. — These  mixtures  are  called  dynamites. 
They  were  introduced  by  Nobel  to  lessen  the  sensitiveness  of  nitro- 
glycerine and  at  the  same  time  retain  its  other  good  qualities.  The 
absorber  of  the  "oil"  is  called  the  "dope,"  which  may  be  selected  to  be 
either  inert  or  active  in  the  explosion. 

The  freezing  temperature  of  all  dynamite  is  that  of  nitre-glycerine,  as 
is  also  its  behavior  when  frozen  and  its  method  for  being  safety  thawed. 
Dynamite  that  does  not  leak  nitro-glycerine  under  the  conditions  under 
which  it  is  to  be  used  is  one  of  the  safest  explosives  known.  It  should 
not  be  shipped,  however,  in  rigid  metallic  cases,  which  accentuate  shocks 
and  vibrations,  but  in  wooden  boxes  in  paper  cartridges  packed  in  saw- 
dust. Thus  packed,  it  has  failed  to  explode  when  dropped  on  the  rocks 
from  a  considerable  height  or  when  struck  by  heavy  weights. 

Dynamite  can  be  heated  with  less  danger  than  nitro-glycerine.  If 
set  on  fire,  it  will  usually  burn  quietly  unless  unfavorable  conditions  are 
present.  If  the  dynamite  is  in  a  closed  box,  its  smoke  cannot  escape  and 
consequently  the  pressure  may  be  raised  enough  to  cause  an  explosion. 
If  caps  or  gunpowder  are  present,  the  fire  will  explode  them  and  the 
resultant  shock  will  detonate  the  dynamite,  If  the  heat  from  the  fire 
causes  the  "oil"  to  exude  from  the  cartilages,  this  "oil,"  if  under  a  static 
head,  will  explode  when  ignited,  as  explained  above.  Again,  the  heat 
from  the  burning  dynamite  may  heat  the  adjoining  unlighted  cartridges 
to  the  exploding  temperature  of  380°  F.  before  they  get  sufficiently 
exposed  to  the  air  to  ignite.  Heated  gradually  in  the  open  so  much  of 
the  "  oil "  may  be  evaporated  that  a  mere  whiff  ensues  when  the  exploding 
temperature  is  finally  reached. 

In  spite  of  all  these  dangerous  contingencies,  several  instances  are  on 
record  where  several  tons  of  dynamite  have  burned  in  conflagrations 
without  exploding.  If  afire  in  cartridges,  it  burns  slowly  like  sulphur, 
but  if  loose  it  will  burn  quickly  like  chaff. 

The  dope  first  used  was  inert  infusorial  earth  or  kieselguhr,  which  will 
safely  absorb  three  times  its  weight  of  nitre-glycerine.  The  resulting 
kieselguhr  dynamite  when  strongest  contains  75  per  cent.  "  oil. "  It  is  a 
pasty,  plastic,  unctuous,  odorless  mass  of  a  yellowish  color  with  a  specific 


14  MINING    WITHOUT    TIMBER 

gravity  of  1.4.  The  effect  of  the  dope  is  to  cushion  the  "oil"  so  that  the 
shock  to  explode  it  must  be  stronger  as  the  percentage  of  dope  becomes 
greater.  It  is  not  possible  to  explode  kieselguhr  dynamites  which  con- 
tain under  40  per  cent,  of  "oil"  and  even  with  60  per  cent,  it  takes  a 
strong  cap. 

The  disadvantage  of  75  per  cent,  dynamite  is  the  exudation  of  "oil" 
on  a  warm  day  or  under  water  so  that  dangers  may  arise  from  having  to 
deal  with  the  sensitive  "oil"  before  suspecting  its  presence.  It  is  thus 
ordinarily  unsafe  to  ship  or  use  and  the  60  per  cent,  strength  is  now 
commonly  sold  as  No.  1.  The  strength  of  kieselguhr  dynamite  is  almost 
equal  to  that  of  its  contained  "oil." 

The  active-dope  dynamites  have  no  such  narrow  limitations  as  the 
inert  types  and  not  only  may  numerous  absorbers  be  used,  but  the  per- 
centage of  nitro-glycerine  may  vary  from  4  to  70  per  cent.  These  explo- 
sives go  under  various  names.  The  common  active  absorbents  are  such 
combustibles  as  wood  meal  or  fiber,  rosin,  pitch,  sugar,  coal,  charcoal,  or 
sulphur,  and  such  oxidizers  as  the  alkaline  nitrates  or  chlorates.  The 
chemical  composition  of  the  oil-dope  mixture  should  be  such  as  to  give 
only  completely  oxidized  products  on  combustion.  The  strength  of  this 
type  is  equal  to  that  of  the  "oil"  plus  that  of  the  explosive  dope  when 
completely  burned.  In  other  words,  black  powder  mixed  with  enough 
"oil"  to  detonate  it  would  all  burn  as  shown  by  the  reaction  of  equation 
(7),  thus  giving  several  times  more  power  than  when  ignited  alone.  The 
density  and  appearance,  as  well  as  the  necessary  strength,  varies  with 
the  dope  and  the  percentage  of  "  oil. "  The  commercial  method  of  rating 
dynamite,  by  its  percentage  of  "  oil,"  is  misleading  as  no  account  is  taken 
of  the  varying  strength  of  the  explosive  dopes. 

Nitro-gelatin  Group. — A  mixture  of  this  group  is  called  a  gelatin 
dynamite.  Somewhat  more  expensive  than  nitro-glycerine,  it  is  prefer- 
able wherever  the  highest  power  is  desired  and,  being  unaffected  by 
water,  it  is  the  best  powder  for  subaqueous  use.  It  is  more  plastic  and 
less  sensitive  than  common  dynamite  and  therefore  easier  to  load  and 
safer  to  transport,  but  it  requires  a  stronger  cap  for  exploding.  The 
military  powder  gelignite,  a  favorite  in  England  and  Japan,  and  forcite 
come  under  this  group. 

Fulminate  Group. — For  percussion-cap  filling,  mercuric  fulminate 
is  mixed  with  a  sufficient  amount  of  some  oxidizer  to  insure  complete 
combustion  on  exploding.  Alkaline-nitrate  oxidizers  may  be  used  but 
potassium  chlorate  is  the  favorite.  The  latter  gives  the  following 
exploding  reaction: 

3Hg(CNO)2  +  2KC103  =  3Hg  +  2KCl+6C02  +  6N.  (18) 

Equation  (18)  shows  that  potassium  chlorate  should  form  22  per 
cent,  by  weight  of  the  mixture,  which  also  contains  a  little  gum  to  give 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  15 

coherence.     Caps  are  designated  by  numbers  or  letters  according  to  the 
amount  of  fulminate  contained.     The  common  series  is. 

Hg  (CNO)2 
Cap  No.  Grains. 

1  4.5 

2  6.0 

3  8.0 

4  10.0 

5  12.0 

6  15.0 
6.5  19.0 

7  23.0 

8  30.9 

The  larger  the  cap,  the  more  expensive,  but  if  the  cap  selected  is  too 
small  to  insure  perfect  detonation  of  the  explosive,  incomplete  combustion 
will  ensue  with  noxious  fumes  and  loss  of  power.  In  general,  dynamite 
requires  stronger  caps  as  the  percentage  of  nitro-glycerine  or  the  temper- 
ature decreases. 

Nitro-benzol  Group. — Although  nitro-benzol  contains  nitryl  it  does 
not  contain  sufficient  oxygen  to  be  an  explosive  and,  when  unmixed  with 
its  oxidizer,  it  can  be  shipped  as  an  ordinary  chemical.  On  this  account, 
the  nitro-benzol  or  Sprengel  group  is  especially  adapted  for  use  in  isolated 
places  far  from  dynamite  factories.  The  favorite  Sprengel  explosive 
is  rackarock,  which  is  a  mixture  of  nitro-benzol  with  the  chlorate  or 
nitrate  of  potassium  or  with  sodium  nitrate,  as  an  oxidizer.  By  mixing 
77.6  per  cent,  of  mononitro-benzol  with  22.4  per  cent,  of  sodium  nitrate, 
we  can  get  the  following  reaction  on  detonation: 

2C6H5(N02)  +  10NaNO3  =  12CO2  +  6N2  +  5H2O +5Na2O.         (19) 

From  equation  (19)  the  percentage  composition,  by  volume,  of  the 
gaseous  product  is, 

C02=   52.2 

N=   26.1 

HaO=  21.7 

100.0 

By  calculation,  the  theoretical  temperature  is  5300°  C.  and  the 
pressure  is  13,800  atmospheres.  Unlike  ignited  black  powder,  rackarock, 
when  properly  detonated,  follows  closely  its  theoretical  reaction  which 
shows  harmless  gases  and  a  temperature  of  79  per  cent,  and  a  pressure  of 
47  per  cent,  of  the  figures  for  nitro-glycerine.  For  practical  use,  the 
oxidizer  of  rackarock  is  handled  alone  in  wax-paper  cartridges  and  the 
required  quantity  of  nitro-benzol  is  not  poured  into  a  cartridge  until 
just  before  charging  the  drill  hole. 


16  MINING    WITHOUT    TIMBER 

Detonating  Permissibles. — These  explosives  practically  all  contain 
either  nitre-glycerine,  nitro-gelatin,  nnitro-bezol  or  ammonium  nitrate 
as  the  detonated  ingredient  and  some  contain  two  or  more  of  them. 
Their  exact  composition  is  usually  kept  secret  by  the  manufacturers, 
but  they  must  pass  the  government  tests  for  temperature  and  flame. 
These  explosives  are  made  of  various  strengths  and  require  stronger  caps 
than  common  dynamites.  Detonation  means  a  quick  generation  of  a 
email  quantity  of  hot  gas  while  the  ignition  of  black  powder  means  the 
slow  production  of  a  large  quantity  of  impure  gases  and  vapors.  A  large 
quantity  of  fine,  unstable  salt  like  magnesium  carbonate,  of  a  steam- 
generating  salt  like  ammonium  nitrate,  or  of  a  substance  with  much 
hygroscopic  moisture  like  wood  meal,  are  the  ingredients  relied  upon  to 
cool  the  quick  small  flame  of  permissibles.  The  compositions  of  a  few 
typical  permissibles  are  as  follows: 

Name.                                Nitro-benzol.  NH<NO|              Ground  Wood.         Water. 

Amvis 4.50  90.0  5.0  0.50 

Ammonite 12.00  88.0 

Electronite 19.0  Ba(NOj)2. .  73.0  7.5  0.50 

Westfalit,    No.    1.    4.5  (rosin)3...  95.0  ...  0.50 

Bellite.  No.  3 5.25  94.0  ...  0.75 

Carbonite 25.0("  oil") . . .  0.50(soda)       34.0  (NaNOa)  40.5 

MISFIRES 

The  cause  of  misfire  depends  upon  both  explosive  and  the  manner 
of  firing.  The  three  classes  of  explosives  with  their  methods  of  firing 
will  now  be  considered. 

Mechanical  Powders  of  Class  I. — In  breaking  coal  with  igniting  pow- 
ders, it  is  inadvisable  to  attempt  to  use  a  missed  hole  if  the  tamping 
must  first  be  dug  out,  therefore  a  new  hole  is  bored,  charged  and  fired 
alongside  the  first.  In  rock  breaking,  where  boring  holes  is  expensive, 
the  tamping  may  be  dug  out  safely  if  only  copper  tools  are  used  when 
approaching  the  powder.  However,  if  the  explosives  are  well  selected, 
and  kept  dry,  and  care  is  taken  in  locating  and  loading  the  holes,  misfires 
will  seldom  occur. 

With  squib-ignition  misfires  may  be  caused  by  (a)  wetness  of  powder; 
(6)  dampness  of  squib;  (c)  loss  of  powder  from  squib;  (d)  squib-hole 
clogged  by  dirt;  (e)  hole  too  long  for  squib  to  recoil  and  reach  powder. 

With  fuse-ignition  misfires  may  be  due  to  (a)  damp  powder;  (6)  cut- 
ting of  fuse  in  tamping;  (c)  imperfect  fuse;  (d)  damp  fuse;  (e)  loss  of 
powder  from  end  of  fuse. 

With  ignition  by  electric  igniter  misfires  may  occur  from  (a)  imperfect 
igniter;  (fe)  damp  igniter;  (c)  wire  broken  in  tamping;  (d)  circuit  im- 
perfectly wired;  (e)  current  leakage  from  poor  insulation;  (f)  current 
deficiency  from  imperfect  or  overloaded  blasting  machine.  The  com- 


EXPLOSIVES    AND    THEIR    USE   IN    MINING  17 

pleteness  of  the  circuit  can  be  tested  before  the  exploding  by  passing  a 
feeble  current  through  a  galvanometer  placed  in  the  circuit. 

Detonating  Powders  of  Classes  II  and  III. — In  breaking  coal  with 
these  powders  it  is  better,  as  with  igniting  powders,  to  bore  and  load  a 
new  hole  than  to  dig  out  the  tamping  from  a  missed  hole.  In  rock  work, 
it  is  good  practice  to  dig  out  the  tamping  from  a  missed  hole  to  within 
only  half  an  inch  of  the  powder  and  then  insert  a  new  primer  cartridge 
with  detonator  and  retamp.  The  excavation  of  tamping  should  be 
cautiously  done  when  approaching  the  powder  and  care  be  taken  not 
to  strike  the  cap. 

Dynamite  should  not  be  allowed  to  remain  long  before  firing  in  water 
holes,  for  the  water  may  displace  the  "oil"  and  perhaps  cause  a  misfire 
or  the  escape  of  "oil"  into  adjoining  crevices  when  it  may  later  be 
struck  by  a  pick  or  drill  and  explode.  Powder  should  never  be  used 
when  even  partly  frozen,  for  the  thawed  portion  may  explode  alone  and 
leave  the  frozen  residue  in  the  hole  or  blow  it  out  into  the  muck  to  become 
in  either  case  a  source  of  danger  for  the  next  shift  of  miners. 

In  firing  a  round  of  holes  in  sequence,  the  explosion  of  one  hole  may 
blow  off  the  primer  of  an  adjoining  hole  whose  remaining  charge  is  there- 
fore left  unexploded  in  the  hole-stump.  Except  for  the  last  contingency, 
and  that  of  two  holes  exploding  simultaneously,  the  counting  of  the  ex- 
ploding reports  gives  a  check  on  detonating  in  sequence  which  is  lacking 
in  simultaneously  firing  by  electricity.  An  electric  cap  may  be  damp  and 
conduct  the  current  through  the  circuit,  without  exploding  itself,  and  a 
missed  hole  will  thus  result.  A  fuse  may  have  a  broken  thread  of  powder 
whose  wrapping  may  catch  fire  and  smoulder  some  time  before  igniting 
the  powder  beyond  the  break.  For  all  these  reasons  the  stumps  of 
blasted  holes  should  be  carefully  examined  before  resuming  work,  and 
where  misfires  are  suspected  a  half-hour  interval  should  elapse  before 
revisiting  the  broken  face. 

Fuse  and  cap  detonation  has  the  last  four  causes  of  misfires  already 
given  for  fuse  ignition  and,  in  addition,  is  liable  to  failure  of  the  cap, 
either  from  dampness,  imperfection,  or  insufficient  strength  for  the  given 
explosive. 

The  causes  of  misfires  already  given  for  electric  ignition  can  be  made 
to  read  correctly  as  the  causes  -with  electric  detonation  by  simply  substi- 
tuting the  word  cap  for  igniter  and  adding  the  requirement  that  the  cap 
must  be  of  adequate  strength. 


CHAPTER  II 
PRINCIPLES  OF  BLASTING  GROUND 

It  is  only  in  recent  years  that  engineers  have  had  much  to  do  with 
the  details  of  underground  excavation,  as  it  was  thought  that  all  the 
schooling  necessary  for  the  successful  miner  could  be  gained  by  practice 
with  a  drill  and  shovel.  It  is  evident,  however,  that  where  rock  breaking 
forms  such  an  important  item  of  expense  as  it  does  in  most  mines,  it 
will  well  repay  study  to  ascertain  if  science  cannot  duplicate  here  the 
same  success  it  has  gained  over  empiricism  in  other  departments. 

After  an  explosion  of  powder  in  the  bore  hole,  Fig.  3,  the  sudden  ex- 
pansion of  the  resulting  gases  will  exert  its  force  equally  in  all  directions 
on  the  bore  hole,  until  either  the  enclosing  rock  or  the  tamping  yields 


Section  Plan 

FIQ.  3. — Cones  of  blasting  rock. 

and  the  gases  escape.  The  rock  will  yield  along  what  is  called  the  line 
of  least  resistance,  which  would  be  be  in  the  assumed  homogenous  rock 
of  Fig.  3.  It  is  evident  that  the  angle  0,  which  the  hole  ab  makes  with 
the  exposed  surface  or  the  free  face  of  the  rock,  can  vary  from  nothing 
to  90  deg.  At  0  =  0  deg.,  there  would  be  no  hole  and  at  90  deg.  the  hole 
would  be  in  the  position  be,  the  line  of  least  resistance,  and  would  give  a 
blown-out  shot.  The  quantity  of  rock  thrown  out  by  the  explosion  would 
have  the  volume  of  a  cone  with  an  altidue  be  or  h,  and  a  base  with  a  radius 
ac,  whose  volume  v  =  l/3  ^r(ac)2  and  where  0  =  45  deg.  (the  usual  con- 
dition for  the  maximum  volume)  ac  =  h  and  we  have  v  —  •" -  —  =  (nearly)  h3, 

3 

or  if  m  is  a  constant,  depending  on  rock,  then  v  =  mh3. 

18 


PRINCIPLES    OF    BLASTING    GROUND  .  19 

For  a  case  with  two  free  rock  faces  if  the  powder  charge  be  placed  at 
e,  with  the  lines  of  least  resistance  eg  and  em  of  equal  length,  the  ex- 
plosion 'will  break  out  two  cones  def  and  fek,  or  nearly  double  the  volume 
for  one  free  face,  so  that  v  =  2mh3.  It  is  similar  for  three  or  more  free 
faces,  so  that  as  a  general  equation  we  have,  if  n  =  the  number  of  free 
faces,  v  =  nmh3. 

From  this  formula  it  can  be  seen  that  a  system  of  mining  should 
be  adopted  which  utilizes  as  many  free  faces  as  possible  in  breaking. 
In  development  work  for  vertical,  horizontal  or  inclined  drives  or  pas- 
sages, we  start  each  round  of  holes  with  one  free  face  and  with  our  cut 
holes  break  out  either  a  cone  or  a  wedge  whose  surface  forms  another 
free  face  for  the  benefit  of  the  other  holes  of  the  round.  In  stoping 
work,  which  must  be  started  from  a  drive,  we  can  always  manage  to 
maintain  two  and  often  three  free  faces  in  homogeneous  rock,  and  in 
stratified  formations  sometimes  four  or  more  faces,  as  a  bedding  plane 
is  often  nearly  the  equivalent  of  a  free  face. 

In  stratified  formations  the  correct  principles  of  breaking  are  especi- 
ally important  for  economy's  sake.  The  simpliest  case  is -that  of  beds 
2  to  4  ft.  thick.  Here  the  holes  should  be  drilled  in  a  plane  parallel 
to  the  beds  because  it  is  evident  that  we  can  more  easily  separate  two 
wet  coins  on  a  table  by  sliding  one  sideways  than  by  trying  to  lift  it 
off  vertically.  Also  these  parallel  holes  do  not  weaken  the  blast  by 
allowing  the  powder  gases  to  escape  through  the  bedding  seam.  Where 
the  beds  are  thin,  say  under  8  in.,  we  encounter  the  possibility,  with 
holes  parallel  to  the  bedding,  of  having  only  the  small  bed  blown  out  that 
contains  the  hole.  For  this  reason  it  is  advisable  to  first  make  a  cut 
by  driving  holes  across  the  bedding  planes  and  then  break  to  the  cut 
with  the  balance  of  the  holes  drilled  parallel  to  the  bedding  plane,  but 
which  now  exert  their  maximum  force  perpendicular  instead  of  parallel 
to  the  beds. 

The  method  of  firing  also  affects  the  pointing  and  the  necessary 
number  of  holes  to  drill  for  breaking.  There  is  a  great  advantage  in 
simultaneous  or  electric  firing  wherever  a  weak  roof  or  the  greater 
danger  from  misfires  with  unskilled  miners  do  not  militate  against  it. 
In  Fig.  3  it  is  evident  that  only  the  cone  abd  and  the  double  cone  dck 
would  be  broken  out  by  the  charges  at  b  and  e  fired  separately,  but  if 
b  and  e  are  not  too  far  apart  and  are  fired  together  the  line  of  detach- 
ment will  be  along  the  lines  abek  instead  of  abdek  and  the  extra  volume  bde 
will  be  broken  with  no  extra  powder  or  drilling.  In  any  case  of  breaking, 
the  pressure  p  produced  by  the  explosive  multiplied  by  the  area  of  its 
section  a  (taken  along  the  axis  of  the  hole)  must  equal  the  ultimate 
tensile  or  shearing  strength  T  of  the  rock  multiplied  by  the  area  of  its 
surface  of  fracture  S  or  pa  =  TS. 

If  pa  is  greater  than  TS  it  means  an  excess  of  explosive  over  that 


20 


MINING    WITHOUT   TIMBER 


required  for  detaching  the  burden.  This  excess  causes  a  "windy" 
shot,  resulting  in  a  greater  air  blast,  a  louder  report  and  a  longer,  hotter 
flame  than  from  a  normal  shot.  A  normal  charge  leaves  traces  of  the 
drill  hole,  but  an  insufficient  charge  leaves  "candlesticks"  in  the  rock 
and  loose  pieces  of  the  burden  have  to  be  blasted  off. 


UNDERGROUND  DEVELOPMENT 

In  illustrating  we  will  take  the  case  of  driving  horizontal  headings  or 
drifts  as  the  same  principles  of  breaking  apply  equally  well  for  inclined 
and  vertical  shafts  and  raises.  The  practical  difference  in  the  latter 
arises  from  the  setting  of  the  drills  and  the  handling  of  the  muck  and 
the  water,  and  the  fact  that  the  length  of  the  section 
in  shafts  generally  makes  the  central  cut  advisable. 
We  will  also  assume,  to  simplify  the  illustrations,  a 
heading  small  and  soft  enough  to  allow  its  breakage 
by  rounds  of  nine  holes  in  three  rows  of  three  holes 
each,  although  often  nine  holes  are  more  effective  in 
four  rows,  one  of  three  and  the  balance  of  two  holes 
each.  For  larger  headings  with  harder  rock,  the  same 
principles  would  apply,  but  more  holes  must  he  added 
for  breaking  the  round.  On  this  basis  we  will  now 
consider  the  following  six  cases  of  formation. 

Case  I.  Homogeneous  Rock  Free  from  Bedding 
Planes  or  Joints  in  the  Face  of  the  Heading. — Since 
this  formation  breaks  equally  well  in  any  direction, 
the  holes  should  be  placed  for  the  most  convenient 
drilling  and  mucking.  For  setting  the  bar  horizont- 
ally, as  is  usual  where  it  is  desired  to  begin  drilling 
before  the  muck  from  the  last  round  is  cleaned  up, 
the  placing  of  Fig.  4  (a)  is  a  favorite.  Here  the^  ad- 
Justalt)le  arm  is  unnecessary  and  the  first  setting  of 
the  bar  is  at  A  to  drill  the  upper  and  middle  rows 
with  the  machine  above  the  bar.  The  second  setting  of  the  bar  is  at 
B  and  the  machine  is  turned  under  it  for  drilling  the  bottom  row  3 
of  lifters.  The  horizontal  rows  of  holes  are  usually  fired  in  the  order 
1,  2,  3,  Fig.  4  (a). 

The  side  instead  of  the  bottom  cut  is  handiest  if  we  wish  to  set  the 
bar  vertically.  We  first  set  up  at  A,  Fig.  5,  and  drill  row  1,  then  at  B 
with  the  machine  on  one  side  to  drill  row  2  and  on  the  other  to  drill  row 
3.  Here  the  vertical  rows  of  holes  are  fired  in  the  order  1,  2,  3,  Fig. 
5  plan.  In  other  to  keep  the  passage  straight,  the  cut  holes  of  row  1 
will  be  put  for  the  next  round  on  the  opposite  side  to  what  is  shown, 
so  that  the  finished  sides  have  a  zig-zag  appearance,  alternately  right  and 


Sec.  (b) 


©A 


Sec.  (c) 


PRINCIPLES    OF    BLASTING    GROUND 


21 


left  as  shown  in  the  plan  of  Fig.  5.  The  middle  hole  of  vertical  row  1 
points  downward,  like  hole  c,  instead  of  flat-wise,  like  the  balance  of 
horizontal  row  1,  so  as  to  throw  out  a  bottom  cut  and  avoid  a  horizontal 
inclination  to  the  face  too  acute  for  rapid  progress  in  a  narrow  heading. 
For  large  tunnel  headings,  8  ft.  square,  in  hard  homogeneous  rock,  the 
cone  or  "Leyner"  center-cut  system  has  recently  permitted  of  very 


Section 
Fid.  5. — Holes  for  headings  with  vertical  bar. 

fast  driving  in  western  metal  mines.  It  is  especially  adapted  to  the 
water  Leyner  drill  on  account  of  the  many  upper  holes  used  and  the 
fact  that  this  drill  is  short  enough  to  allow  the  sharp  pointing  of  the 
holes  with  two  settings  of  the  bar.  For  hard  steel  ore  and  jasper  in  a 
Michigan  iron  mine,  this  system  was  thus  applied. 

In  Fig.  6,  A  is  the  bar  in  first  position  for  two  machines  and  from  its 


J    J 


Front  Elev.  Section 

FIG.  6. — Holes  for  Leyner  tunnel  cut. 

top  the  four  back  holes,  Nos.  9,  10,  11  and  12,  are  drilled.  The  machines 
are  then  tipped  forward  until  the  crank  can  just  turn  and  clear  the  back 
or  top  of  the  drift  for  drilling  the  top  center  cut  holes  Nos.  1  and  2, 
while  finally  they  are  turned  under  the  bar  for  side  holes  Nos.  5,  6,  7 
and  8.  The  bar  is  then  changed  to  position  B,  the  machines  are  set  up  on 
top  and  side  holes  Nos.  13  and  14  are  drilled.  Then,  after  turning  the 


22  MINING    WITHOUT   TIMBER 

machines  under  the  bar,  they  are  tipped  up  in  front  so  the  crank  just 
clears  the  bottom  of  the  drift  and  holes  Nos.  3  and  4  are  drilled  about  to 
meet  Nos.  1  and  2  in  the  center  of  the  heading.  The  four  lifters,  Nos. 
15,  16,  17  and  18,  are  the  final  holes.  In  softer  and  better-breaking 
ground,  cut  holes  Nos.  5  and  6,  one  lifter  and  one  back  hole  can  be  left 
out,  but  the  four  cut-holes,  Nos.  1,  2,  3  and  4,  are  nearly  always  used 
and  are  pitched  up  and  down  and  in,  to  meet  about  in  the  center. 

The  five  remaining  cases  are  given  for  regularly  stratified  rock,  but 
the  joints  or  cracks  of  massive  rock  may,  like  bedding  planes,  often  be 
utilized  for  breaking. 

Case  II.  Rock  in  Horizontal  Beds;  (a)  Medium  Thick  Beds. — Here 
the  best  results  from  the  powder  can  be  got  by  two  settings  of  the  bar 
vertically  and  following  the  drilling  and  firing  directions  given  above 
for  the  method  illustrated  by  Fig.  5.  In  the  disseminated  lead  mines 
of  southeastern  Missouri  (Example  10,  Chapter  VIII),  this  method  is 
modified  as  follows: 

For  a  drift  10  ft.  wide  by  6  1/2  to  7  ft.  high,  12  to  13  holes  are  needed, 
placed  in  three  rows  horizontally  by  four  rows  vertically.  The  bar  is  set 
up  once  to  drill  each  vertical  row  of  holes,  four  set-ups  being  necessary 
to  complete  a  round.  Each  vertical  row  is  fired  separately  by  fuse  and 
dynamite  and  as  only  three  or  four  holes  are  fired  at  a  time,  not  enough 
smoke  or  broken  rock  is  produced  to  prevent  the  drillers  from  setting 
up  again  very  soon  after  blasting.  This  method  with  three  shifts  of  two 
drill  men  each  allows  an  advance  of  5  to  7  ft.  in  24  hours  with  2  3/4-in. 
drills.  By  the  former  center-cut  system,  two  drills  and  four  men  were 
able  to  advance  only  10  to  15  per  cent,  faster  than  by  the  one  drill  and 
the  side-cut  method  just  described,  all  loading  and  tramming,  in  each 
case,  having  been  done  by  muckers. 

(b)  Thin  Beds. — Here,  as  already  explained,  the  cut-holes  must  cross 
the  bedding  planes.  A  bottom  cut  is  advisable.  The  bar  is  set  horizon- 
tally at  A,  Fig.  4  (6).  Often  all  three  rows  can  be  drilled  direct  although 
sometimes  the  use  of  the  adjustable  arm  on  the  bar  is  necessary  to  get 
the  correct  pointing  of  the  holes.  The  holes  of  row  1  are  fired  first  and 
break  out  the  cut  to  the  bedding  plane  on  the  floor  of  the  heading.  Be- 
fore loading  the  row  of  cut-holes,  it  is  often  helpful  to  stop  up  their  bed- 
ding planes,  around  the  powder,  with  clay  but  this  precaution  is  unneces- 
sary in'  the  two  upper  rows  where  the  holes  are  parallel  to  the  beds. 

Case  III. — Rocks  in  Vertical  Beds  Parallel  to  Heading;  (a)  Medium 
Thick  Beds. — This  case  requires  the  bottom  cut  of  Fig.  4  (a)  which  has 
already  been  described  under  Case  I.  The  use  of  this  method  in  the 
vertical  copper  veins  of  Butte,  Mont.,  is  as  follows:  The  placing  of  holes 
is  shown  in  Fig.  7  for  the  12-hole  system,  although  for  most  rock  nine 
holes  are  ample,  the  center  holes  of  rows  2,  3,  and  4  being  omitted.  For 
this  arrangement  the  drill  bar  (with  adjustable  arm)  need  only  be  set 


PRINCIPLES    OF    BLASTING    GROUND 


23 


up  once  vertically,  as  shown.  The  round  of  holes  is  usually  loaded  and 
fired  at  one  time  and  goes  off  in  the  order  of  1,  2,  3,  4.  Some  of  the 
miners  regulate  the  explosions  by  cutting  the  fuse  of  different  lengths 
and  spitting  them  simultaneously  while  held  together  hi  the  hand,  and 
others  by  cutting  all  the  fuse  of  the  same  length  and  spitting  them 
separately  in  the  required  order. 

(6)  Thin  Beds. — The  solution  of  this  case  follows  Fig.  5  and  also 
resembles  Case  II  (6)  except  that  here  the  side  instead  of  the  bottom  cut 
is  used.  With  one  setting  of  the  bar,  the  three  vertical  rows  K,  2  and  3 
may  be  drilled  and  shot  in  the  same  order,  row  K  breaking  out  the  cut, 
along  a  side  bedding  plane,  mn,  and  rows  2  and  3  breaking  to  the  cut. 


Longit.  Section  Sec.  A-B 

FIG,  7  — Holes  for  sloping. 

Here  it  is  not  so  necessary  for  alignment,  as  in  Case  II  (a),  to  alternate 
the  cut  on  each  side  of  the  heading,  but  it  is  often  an  advantage  especially 
where  the  vertical  bedding  planes  are  ill  denned. 

Case  IV. — Rocks  in  Vertical  Beds  Cutting  the  Heading  at  an  Angle;  (a) 
Medium  Thick  Beds. — If  the  cutting  angle  which  the  bedding  plane 
makes  with  the  side  of  the  heading  is  45  deg.  or  less,  the  method  of 
Fig.  4  (a)  is  usually  preferable.  If  the  cutting  angle  is  more  than  45 
deg.,  the  choice  between  the  methods  of  Fig.  4  (a)  and  of  Fig.  5  will 
often  be  merely  a  question  of  convenience  in  setting  the  bar  horizontally 
or  vertically,  respectively. 

(6)  Thin  Beds. — With  a  cutting  angle  of  45  deg.  or  less  the  method 
of  Fig.  5  is  the  best.  Where  the  cutting  angle  is  more  than  45  deg., 
the  choice  between  the  methods  of  Fig.  4  (a)  and  Fig.  5  depends  on  set- 
ting the  bar  as  in  Case  IV  (a). 


24  MINING    WITHOUT    TIMBER 

Case  V. — Rocks  in  Inclined  Beds  Dipping  Toward  the  Floor  of  the 
Heading. — For  either  medium  thick  or  thin  beds  the  method  of  Fig.  4 
(a)  is  the  best.  Care  must  be  taken,  however,  in  the  case  of  beds  dipping 
over  45  deg.  to  stop  the  holes  of  the  horizontal  row  1  at  the  last  bedding 
plane  which  intersects  the  face  of  the  heading  above  the  floor. 

Case  VI. — Rocks  in  Inclined  Beds  Dipping  Away  from  the  Floor  of  the 
Heading.-— For  either  medium  thick  or  thin  beds  the  method  of  Fig.  4  (c) 
should  be  used.  The  bar  is  set  up  at  A  for  row  2  and  at  B  for  rows  1  and  3. 
The  order  of  firing  the  horizontal  rows  of  holes  is  1,  2  and  finally  3. 
The  end  of  the  holes  in  row  1  should  be  stopped  beneath  the  last  bed- 
ding plane  intersecting  the  face  of  the  tunnel  under  the  roof  in  order 
to  utilize  this  plane  as  a  free  face  in  breaking. 

SURFACE  EXCAVATION  AND  UNDERGROUND  STOPING 

In  some  kinds  of  deposits,  especially  the  huge  copper-bearing  por- 
phyry lenses  and  the  Lake  Superior  iron  mines,  much  time  is  often 
saved  by  drilling  all  the  holes  possible  in  the  periphery  of  a  heading  in  the 
ore  from  the  same  set-ups  that  are  used  in  drilling  the  face.  These 
peripheral  holes  can  then  be  left  untouched  until  the  stoping  of  that 
section  begins,  when  they  can  be  easily  loaded  and  fired. 

Holes  for  stoping  may  be  placed  according  to  the  direction  in  three 
groups,  (1)  down  holes  (2)  flat-holes,  and  (3)  uppers.  A  dip  of  about 
45  deg.  downward  and  upward  can  be  assumed  to  make  the  limit 
between  groups  (1)  and  (2)  and  of  (2)  and  (3)  respectively,  although  the 
division  between  (2)  and  (3)  is  really  marked  by  the  angle  of  repose  of  the 
cuttings,  that  is,  when  the  hole  becomes  self-cleaning,  which  may  often 
mean  a  steeper  dip  than  45  deg.  The  speed  of  cutting  with  recipro- 
cating drills  depends  on  the  removal  of  cuttings  after  each  stroke  to 
expose  a  fresh  face.  Therefore  with  these  drills  down  holes  drill  easiest, 
then  uppers,  and  lastly  flats.  Using  the  hammer  drills  with  hollow  bits 
cleaned  by  water  or  air-jets,  there  is  less  difference  in  drilling  speed 
for  different  directions  of  pointing. 

Down  Holes;  Underground. — Down  holes  are  used  underground  in  the 
underhand  benches  of  tunnels  or  metal  mines.  To  start  this  system,  a 
heading  ah,  Fig.  8,  is  run  at  the  top  of  the  tunnel  or  stope  and  the  down 
holes  put  in  its  floor  for  the  first  bench.  The  depth  of  this  bench  is 
limited  by  the  length  of  the  bit  which  can  be  inserted  in  the  hole  and  that 
depends  on  the  height  of  the  heading  which  is  usually  around  7  ft.  so 
that  the  ordinary  railroad  tunnel,  20  to  25  ft.  high,  requires  two  benches 
and  two  settings  of  the  tripod  at  a  and  b,  Fig.  8  to  reach  the  bottom. 
These  bench  holes  point  downward  anyhow  but  often  an  advantage 
may  be  taken  of  the  structure.  Thus  with  horizontal  beds,  the  holes 
of  the  first  bench  can  be  terminated  at  a  bedding  plane  which  the  gases 


PRINCIPLES    OF    BLASTING    GROUND 


25 


from  the  explosion  will  enter  and  thus  exert  a  lifting  action  on  the  mass 
to  be  broken  off. 

Where  there  is  a  choice  of  plans,  a  heading  can  often  be  given  in  a 
direction  that  will  take  the  maximum  advantage  of  the  bedding  and 
joint  planes  for  breaking,  both  in  driving  and  stoping.  On  this  principle, 
the  rooms  of  coal  mines  are  usually  laid  out  perpendicular  to  the  line 
of  the  main  joint  planes  of  the  coal  seam  or  to  the  "face  cleat." 

Down  Holes;  at  Surface.  —  Above  ground  the  only  limit  to  the  depth 
of  the  hole  is  the  capacity  of  the  drill.  In  considering  breaking  from 
deep  holes  we  have  a  choice  of  two  methods  (a)  multi-charging,  (b) 
chambering. 

In  the  drill  hole  ab  of  Fig.  9,  it  is  evident  that  a  charge  of  explosive 
at  any  point  b  will  only  break  out  a  cone  like  cbd  where  eb  is  the  line  of 


ha 


Fia.    8. — Holes  for  underhand  stope. 


FIG.  9.— Holes  for  high  bench. 


least  resistance.  In  order  to  break  the  whole  length  ab  by  multi-charg- 
ing, other  charges  of  explosives  as  /  and  g  would  be  placed  along  the  hole, 
with  tamping  between,  and  all  be  set  off  by  simultaneous  firing.  In  this 
way  the  whole  mass  abd  would  be  detached. 

By  chambering,  the  breaking  from  a  long  hole  would  be  achieved 
differently.  Instead  of  the  hole  being  placed  near  the  face  hd  of  the 
bench  as  is  the  hole  ab  (because  of  its  small  section  for  developing 
explosive  pressure),  the  hole  mn  would  be  placed  back  from  the  face  so 
that  nc,  the  line  of  least  resistance  in  homogeneous  rock,  would  be  only 
a  little  shorter  than  the  length  of  the  hole  above  the  chamber  at  n.  The 
chambering  is  effected  by  shattering  the  bottom  of  the  hole  with  high- 
power  dynamite  so  that  the  final  shape  of  the  chamber  approaches  a 
sphere.  In  France  this  chambering,  in  limestone,  is  performed  with 
hydrochloric  acid,  each  dose  of  neutralized  acid  being  washed  out  and  a 
new  one  poured  in  until  the  chamber  is  of  the  required  size.  When  the 
chamber  is  filled  with  gunpowder  or  low-power  dynamite  and  exploded, 


20  MINING    WITHOUT   TIMBER 

it  will  exert  nearly  as  much  force  upward  as  horizontally  and  will  break 
out  a  mass  along  the  surface  of  fracture  qnp. 

The  choice  between  multi-charging  and  chambering  depends  on  the 
varying  conditions  of  formation,  drilling  and  exploding.  In  a  fissured 
formation,  chambering  has  often  an  advantage  because  the  explosive 
maybe  localized  in  a  solid  portion  of  the  rock,  although  it  often  needs  the 
use  of  two  kinds  of  explosives,  one  for  chambering  and  the  other  for 
breaking.  Where  it  is  desired  to  break  off  only  a  thin  slice  like  hab,  Fig. 
4,  from  the  cliff,  it  is  evident  that  multi-charging  should  be  resorted  to. 
When  an  even  topography  will  allow  the  handling  of  the  portable  steam 
or  electric  churn  drill  for  a  3-in.  to  12-in.  hole  (instead  of  the  reciprocating 
drill  for  a  1  1/2-in.  hole),  the  multi-charging  method  will  permit  the 
drilling  and  breaking  of  a  much  longer  hole  than  would  be  feasible  by 
chambering. 

Flat  Holes;  Underground. — Of  the  three  groups,  flat  holes  are  the 
most  difficult  to  drill,  especially  those  which  are  pointed  above  the 
horizontal  for  the  reason  that  they  neither  hold  water  or  discharge  their 
cuttings  by  gravity.  This  group  is  much  used  in  the  overhead  stoping 
system  with  piston  drills  as  the  drill  tripod  can  be  set  on  the  broken 
rock.  Thus  flat  water  holes  which  are  easier  to  load  than  uppers  and 
free  from  their  dust  can  be  drilled  at  a  fair  speed.  In  overhand  stoping 
with  a  weak  back,  as  in  the  vertical  veins  of  Butte,  Mont.,  flat  holes 
have  also  an  advantage  over  uppers  as  the  timber  sets  can  be  carried  next 
to  the  back  and  the  drilling  can  proceed  under  the  lagging.  Thus  in 
Fig.  7  at  C  rows  1  and  2  are  water  holes  and  only  row  3  need  be  drilled  dry. 

In  the  zinc  district  of  Joplin,  Mo.,  flat  holes  are  used  instead  of  the 
usual  down  holes  to  break  the  benches  below  the  heading  of  the  under- 
hand stoping  system  as  described  under  Example  10  of  Chapter  VIII. 


*\        *\         -»° 

T 

£ 

t  /  '«?  ''  /a.                   c\  \  hx    a 

H 

C 

y 

_r 

Fia.   10. — Holes  for  seam. 


In  driving  coal  headings  or  rooms  by  "blasting  off  the  solid,"  flat 
holes  bored  by  augers  are  generally  used  and  are  placed  similarly  to 
those  shown  for  headings  in  flatly  bedded  rock  in  Fig.  5.  In  the  location 
of  the  horizontal  rows  of  holes,  the  character  of  the  bedding  planes  be- 
tween the  coal  seam  and  its  roof  and  floor  must  be  considered.  If  the 
roof  is  "tight,"  the  shot  must  exert  a  strong  shearing  force  to  separate  it. 
This  is  achieved  by  slanting  the  row  of  holes  sharply  upward  and  ter- 


PRINCIPLES    OF    BLASTING    GROUND 


27 


minating  them  at  the  tight  plane.  A  similar  remedy  is  applied  to  a  tight 
floor.  In  many  seams  the  coal  is  cut  up  into  cubes  by  two  sets  of  joint- 
planes  perpendicular  to  the  bedding  planes,  called  the  "face"  and  "end" 
cleats,  which  condition  makes  breaking  easy. 

The  shearing  of  a  coal  face,  before  shooting,  takes  the  place  of  the  cut 
holes  in  blasting  off  the  solid  and  the  smaller  charges  allowable  for  the 
former  method  not  only  save  explosive  but  prevent  the  shattering  of  the 
roof.  With  coal  sheared  vertically  along  one  rib  of  a  heading,  the  holes 
for  breaking  would  be  placed  like  vertical  rows  2  and  3,  Fig.  5.  Where 
the  shear  is  made  horizontally  as  in  the  undercut  xy,  Fig.  10,  it  is  custo- 
mary in  a  thick  seam  of  coal  to  place  the  first  or  "buster"  shot  at  6  in 
order  to  break  out  the  triangular  prism  of  coal  abc.  Then  when  the 
shattered  strip  gfh  has  been  removed  by  the  pick,  we  have  dm  and  en 
instead  of  dt  and  cs  for  the  line  of  least  resistance  from  the  corner  holes 
d  and  e,  by  which  last  the  balance  of  the  undercut  coal  can  now  be  easily 
shot  down.  For  a  thin  vein  of  coal,  the  "buster"  shot  would  be  located 
at  K  on  a  level  with  the  corner  holes  and  it  would 
break  out  the  triangular  prism  tKs  as  thick  as  the 
seam. 

The  undercut  shown  in  Fig.  10  is  that  made  by 
a  hand  or  power  pick.  Being  a  height  of  12  in.  or 
so  in  front  with  a  downward  slope  to  4  in.  in  the 
back,  its  shape  allows  the  "buster"  shot  to  throw 
much  of  the  coal  out  of  the  undercut,  so  that  the 
strip  gfh  can  be  easily  extracted  by  the  pick  to 
prepare  for  the  corner  shots.  When  the  undercut, 
however,  is  made  by  a  chain  machine,  it  is  of  uni- 
form height  of  only  about  4  in.,  and  the  "buster" 
shot  may  not  throw  the  coal  outward.  It  is  then 
often  advisable  to  place  an  extra  "snubbing"  shot 
at  /  to  flatten  down  the  detached  prism  abc  so  that 
the  shots  d  and  c  can  be  made  effective  without  first 
cleaning  out  the  broken  coal  underneath. 

Flat  Hole;  Surface. — In  loosening  huge  banks  of  placer  gravel  in 
California  before  hydraulicking,  small  adits  have  been  used  with  cross- 
cuts at  their  ends  to  hold  the  explosive.  From  a  breaking  stand-point, 
these  adits  correspond  to  flat  drill-holes  with  chambered  ends.  The 
same  method  has  also  been  employed  for  breaking  great  masses  of  rock 
in  quarries  or  excavations.  Often  a  shaft  has  been  sunk  as  an  entrance  to 
the  explosive  chamber  instead  of  an  adit.  Sometimes  two  cross-cuts 
from  the  adit  may  be  made  for  explosive  chambers,  as  shown  in  Fig.  11. 
There  only  the  crosscuts  cd  and  ab  would  be  packed  with  gunpowder  or 
low-power  dynamite,  while  the  adit  itself  would  be  blocked  with  timber 
or  masonry  bulkheads  wherever  it  met  the  crosscuts.  Elsewhere  it  would 


11.— Blasting  by 
tunnels. 


28  MINING    WITHOUT   TIMBER 

be  packed  with  sand.  For  firing,  electric  fuzes  or  caps  would  be  placed 
in  the  explosive  at  intervals  of  about  10  ft.  Finally  they  would  all  be 
connected  by  wiring  in  order  that  they  might  be  fired  simultaneously  by 
electricity  clk  being  the  line  of  least  resistance.  The  chamber  cd  would 
break  out  the  cone  gclfl,  and  the  chamber  ab  would  break  out  the  prism 
halclg,  the  plan  of  the  line  of  fracture  being  mabn. 

The  same  breaking  equation,  pa  =  TS,  applies  as  in  the  case  of  drill 
holes,  the  factor  a  being  the  area  of  the  cross  section  of  the  explosive 
taken  along  the  axis  of  the  crosscut. 

Uppers. — Uppers  are  seldom  used  on  the  surface  but  are  common  in 
underground  work  not  only  in  tunnel  headings  and  raises,  but  also  in 
overhand  stoping.  In  excavating  overhand  stopes  with  square-set 
timbering,  it  is  sometimes  more  efficient  to  drill  the  back  with  up.pers 
as  at  B,  Fig.  7,  instead  of  the  flats  at  C  used  in  Butte  practice.  In  the 
great  stopes  of  the  Portland  mine,  at  Cripple  Creek,  Colo.,  where  the  pay 
shoot  was  in  places  120  ft.  wide  and  400  ft.  long,  the  ore  hard  and  the 
back  strong  enough  to  stay  up  across  the  vein  for  several  sets  ahead  of 
the  timbermen,  it  was  found  that  the  fastest  breaking  was  accomplished 
by  drilling  uppers  from  piston  drills  set  on  tripods,  one  drill  being  used 
in  every  set  across  the  stope. 


CHAPTER  III 
COMPRESSED  AIR  FOR  MINING 

In  drilling  with  piston  rock  drills  a  high  pressure  gives  a  stronger 
withdrawing  force  on  the  bit  which  tends  to  prevent  sticking  in  fissured 
ground  and  thus  greatly  increases  the  speed  of  boring.  In  hard,  tough 
ground,  like  specular  hematite  or  certain  intrusives,  a  high  air  pressure  is 
necessary,  if  it  is  desired  to  strike  a  blow,  severe  enough  to  cut  the  rock, 
with  a  light  portable  machine.  In  a  certain  mine,  using  40  drills  in  hard 
and  fissured  ground  the  rock  broken  per  machine  was  increased  about  20 
per  cent,  by  the  simple  expedient  of  advancing  the  air  pressure  from  75 
to  100  pounds.  A  low  pressure  system  requires  larger  pipes  to  deliver 
the  same  power  and  heavier  pumps  and  hoists  in  the  mine  to  accomplish 
a  given  amount  of  work  than  an  equivalent  equipment  working  under 
high  pressure. 

The  economical  limit  of  pressure  depends  in  a  given  case  on  commer- 
cial considerations,  costs  of  fuel,  labor  and  supplies,  which  in  turn  are 
governed  in  considerable  degree  by  the  mechanical  efficiency  of  the  plant. 
The  high  pressure  limit,  except  for  haulage  purposes  is  about  120  pounds. 


FIG.  12. — Relations  of  volume  and  pressure  in  air  compression. 

It  is  wasteful  to  heat  the  air  during  compression  to  a  higher  tempera- 
ture than  that  of  the  mine,  as  radiation  in  the  pipe-line  will  cool  any 
warmer  air  before  it  reaches  the  motor.  A  proof  of  this  statement 
follows :  Let  V  and  P  be  respectively  the  volume  and  pressure  of  free  air 
at  the  beginning  of  compression,  and  in  the  theoretical  indicator  card, 
Fig.  12,  in  which  0  is  the  origin  of  coordinates,  let  the  abscissa  of  point  a 
be  V  and  the  ordinate  be  P.  Let  V  and  P  be  the  volume  and  pressure 
of  air  at  any  point  of  the  stroke,  during  its  compression  by  a  reciprocating 
piston.  Then  if  the  temperature  due  to  the  heat  of  internal  friction  is 

29 


30  MINING    WITHOUT   TIMBER 

retained  in  the  air,  we  have  adiabatic  compression  and  get  the  curve 
a  b  f,  the  equation  of  which  is 


the  value  of  y  being  1.406  for  dry  air  and  somewhat  less  for  the  ordinary 
atmosphere,  and  p  being  the  resultant  pressure  and  v  the  resultant 
volume. 

If  the  temperature  is  kept  constant  during  compression,  by  removing 
the  internal  heat  as  fast  as  generated,  we  have  isothermal  compression 

and  get  the  cruve  acd,  the  equation  for  which  is  p  =  PVl~  J.     Finally, 

the  work  lost  by  cooling  the  air,  from  the  final  adiabatic  temperature  to 
that  of  the  free  air,  is  measured  by  the  area  acdfb,  the  total  work  of 
compression  for  one  stroke  of  the  piston  being  area  afmn. 

THEORY  OP  THE  INTERCOOLER 

Although  isothermal  compression  is  the  ideal,  practical  difficulties 
prevent  its  attainment.  The  air  can  be  cooled  in  the  compression  cylin- 
der by  a  water  spray,  but  this  method  requires  too  slow  a  machine  to 
compete  with  dry  compression  and  external  cooling.  It  can  be  easily 
shown,  mathematically  or  by  an  indicator  card,  that  water-jacketing 
the  compression  cylinder  has  practically  no  effect  in  cooling  the  air, 
although  it  is  useful  in  keeping  the  bearing  surfaces  cool  enough  for 
lubrication. 

In  Fig.  12,  the  adiabatic  and  isothermal  curves  get  farther  apart  as 
the  pressure  increases,  so  that  the  work  lost  by  adiabatic  compression 
increases  at  a  faster  ratio  than  the  pressure.  To  avoid  this  increase  for 
high  pressures,  a  compression  in  two  stages,  with  a  surface  intercooler 
between  the  high-  and  low-pressure  cylinders,  is  frequently  used.  Unfor- 
tunately, few  of  the  standard  machines  have  a  large  enough  intercooler 
to  insure  that  the  compressed  air,  entering  the  high-pressure  cylinder,  is 
as  cool  as  the  free  air  entering  the  low-pressure  cylinder  when  the  machine 
is  running  full  speed.  It  will  aid  the  intercooler,  if  the  free  air  is  sucked 
into  the  low-pressure  cylinder  from  the  coolest  available  place. 

In  the  diagram,  Fig.  12,  K  is  the  pressure  at  which  the  air  leaves  the 
low-pressure  cylinder  to  pass  through  the  intercooler  and  enter  the  high- 
pressure  cylinder.  The  following  cycle  then  takes  place  with  a  perfect 
intercooler.  In  the  low-pressure  cylinder  the  air  is  compressed  adia- 
batically  from  a  to  b,  reduced  in  the  intercooler  to  the  volume  at  point 
c  and  then  compressed  adiabatically  in  the  high-pressure  cylinder  from 
c  to  e,  the  total  work  of  compression  being  the  area  ab  c  emn.  Thus 
the  saving  of  work  by  the  use  of  the  intercooler  is  represented  by  the  area 
cefb,  from  which  must  be  deducted  any  work  expended  in  circulating 


COMPRESSED    AIK    FOR    MINING  31 

the  cooling  water.  In  the  design  of  the  machine,  the  ratio  of  the  diam- 
eters of  the  low-pressure  cylinder  and  the  high-pressure  cylinder  are 
taken  so  that  the  area  a  b  k  n  is  equal  to  area  c  e  m  k  f or  average 
conditions. 

There  need  be  little  difference  in  the  efficiency  of  the  steam  ends 
between  high-  and  low-pressure  compression.  With  a  cross-compound 
air  end,  the  steam  end  can  also  be  compound  and  for  a  single-stage  air 
end  the  machine  can  be  tandem-compound.  The  air-pressure  governor 
has  now  been  perfected  and  for  the  usual  variable  loads  of  mine  work, 
is  indispensable  for  any  pressure,  though  it  requires  a  duplex  machine 
to  avoid  a  stoppage  on  dead  center  with  no  load. 

PREHEATERS 

In  the  case  of  the  air  motor,  the  compression  process  is  reversed. 
The  air  on  entering  the  motor  in  the  mine  has  the  pressure  and  volume 
of  point  d  (Fig.  12)  and  in  a  simple,  unheated  motor  cylinder  will  expand 
adiabatically  along  the  line  d  g  h.  Should  the  air  be  preheated  to  the 
volume  of  point  /  it  will  then  expand  along  the  adiabatic  line  /  b  a  with  a 
gain  of  work,  over  the  unheated  case,  equal  to  area  a  h  df. 

With  two-stage  expansion,  the  air  may  be  preheated  before  entering 
the  low-pressure  cylinder  to  e,  then  expand  adiabatically  to  c,  next  pass 
through  an  interheater  so  as  to  reach  6  on  entering  the  high-pressure 
cylinder  and  finally  expand  adiabatically  to  a.  Heating  during  expan- 
sion, like  cooling  during  compression,  gains  in  its  relative  effect  on  the 
efficiency,  the  higher  the  pressure.  Aside  from  its  gain  in  work,  heating 
is  often  necessary  to  prevent  freezing  of  the  exhaust  when  the  air  is 
damp  and  cold  on  entering  the  motor. 

Owing  to  the  small  size  and  portability  of  rock  drills  preheaters  are 
for  this  service  out  of  place,  but  for  large  hoists  and  pumps,  with  high- 
pressure  air,  they  are  always  to  be  recommended.  In  the  operation  of 
the  preheater  the  compressed  air  passes  through  a  vessel  containing 
heated  tubes  of  sufficient  radiating  surface  for  the  purpose.  These 
tubes  may  be  heated  by  a  coal,  coke  or  oil  fire,  but,  since  smoke 
contaminates  the  atmosphere  of  the  mine,  steam-heating  is  often  both 
convenient  and  economical.  In  an  air  heater  it  is  possible  to  utilize 
steam  more  efficiently  than  in  the  best  condensing  engine,  for  both 
the  latent  and  visible  heat  of  the  steam  are  absorbed  by  the  air 
and  turned  into  work  without  frictional  losses  greater  than  the  motor 
would  suffer  with  unheated  air.  With  steam  heating  the  only  im- 
portant loss  is  that  due  to  radiation  in  the  supply  pipe  from  the 
boilers,  and  by  proper  covering  this  can  be  made  small.  In  the  500-gal. 
Dickson  pumps,  installed  in  the  Anaconda  mines  at  Butte  in  1899,  the 
air  was  successfully  heated  by  steam  in  both  the  preheaters  and  the  inter- 
heaters  for  the  compound  cylinders. 


32  MINING    WITHOUT    TIMBER 

High-pressure  pipe-lines,  though  smaller  in  diameter,  require  more 
care  to  keep  them  tight  than  lines  for  low  pressure,  and  the  velocity  of 
exit  of  air  from  a  leak  varies  directly  as  the  square  of  the  pressure. 

The  loss  of  power  from  the  common  practice  of  blowing  out  powder 
smoke  with  the  air  hose  is  the  greater  the  higher  the  pressure,  for  the 
ventilating  efficiency  depends  only  on  the  quality  of  free  air  discharged. 
With  pipes  properly  proportioned  for  the  quantity  of  air  to  be  delivered 
the  frictional  line  losses  will  be  moderate  with  either  pressure,  if  care  be 
taken  to  avoid  unnecessary  bends  and  to  use  gate  valves  instead  of  globe 
valves. 

The  compressor  should  discharge  its  air  into  a  receiver  the  cooling 
action  of  which  will  not  only  at  cone  reduce  the  volume  to  that  which  it 
will  have  in  the  mine,  but  will  also  precipitate  any  extra  moisture  and 
keep  it  from  entering  the  pipe-lines.  A  good  device  for  the  surface 
receiver  is  a  condemned  boiler,  set  in  a  wooden  tank  in  which  is  water 
circulating  through  the  boiler  tubes,  while  the  compressed  air  fills  the 
shell.  Underground  the  receivers  need  only  be  plain  steel  shells  for 
storage,  but  they  must  be  numerous  and  large  enough  to  preserve  the 
pressure  constant  under  the  variable  power  requirements.  Preheaters 
in  use  serve  as  receivers. 

When  air  is  used  for  haulage  it  needs  a  special  piping  system  to  hold 
the  requisite  pressure  of  1000  Ibs.  upward.  This  piping  also  serves  as  a 
receiver  and  accumulator  of  air  between  locomotive  chargings  so  that 
the  compressors  can  be  run  under  a  constant  load.  It  is  evident 
that  the  piping  system  will  need  a  lesser  proportionate  capacity 
as  receiver  the  greater  the  number  of  locomotives  supplied,  for 
each  charging  will  involve  a  less  relative  displacement  of  air.  Under 
the  usual  traffic  and  air  pressure  a  pipe  line  of  6  to  12-in.  dia.  is  amply 
large,  both  for  distribution  and  storage  of  air,  without  placing  tank 
receivers  at  the  stations. 

The  air  ends  of  compressors  for  haulage  systems  should  be  at  least 
4-stage,  of  moderate  speed  and  with  ample  intercooling  surfaces;  for 
Fig.  12  shows  how  fast  the  power  loss  due  to  inefficient  cooling  increases 
with  the  pressure.  Until  recently  the  locomotives  were  single-stage  and 
had  consequently  a  low  efficiency  and  capacity;  but  the  new  compound, 
Porter  locomotive  obviates  these  troubles  and  gives  air-haulage  a  chance 
for  extension  beyond  its  present  special  field  of  gaseous  or  dusty  coal 
mines. 


CHAPTER  IV 
PRINCIPLES  FOR  CONTROLLING  EXCAVATIONS 

The  art  of  timbering  is  not  synonymous  with  that  of  the  control  of 
ground  as  some  suppose;  a  good  carpenter  can  frame  timber  better  than 
any  miner,  but  unless  he  places  it  underground  as  directed  by  the  latter, 
his  accurately  jointed  sets  are  liable  to  prove  worthless  for  the  purpose 
intended.  The  subject  of  ground  control  naturally  divides  itself  under 
two  topics:  I.  The  control  of  the  roof  of  an  excavation;  II.  The  control 
of  the  sides  and  floor  of  an  excavation  and  of  the  whole  overlying  forma- 
tion. Both  topics  will  be  considered  separately  before  their  inter-relation 
will  be  discussed.  In  practice,  we  have  not  only  to  consider  the 
freshly  broken  surfaces  of  an  excavation,  but  their  future  conditions  after 
exposure  to  the  weathering  action  of  the  mine  atmosphere. 

CONTROL  OF  THE  ROOF 

(a)  Roof  over  a  Horizontal  Room. — This  case  is  the  simplest  and 
occurs  in  mining  horizontal  seams  or  beds.  Let  abb' of ,  Fig.  13,  represent 
the  cross-section  of  a  rectangular  room  excavated  in  a  team  of  the  thick- 


FIG.   13. — Homogeneous  or  horizontally-bedded  roof. 

ness  aa'.  Then  the  support  of  the  roof  over  the  opening  ab  depends  upon 
the  immediately  overlying  formation.  The  structure  of  the  last  falls 
usually  under  one  of  the  five  following  cases:  (1)  homogeneous,  (2) 
horizontally  bedded,  (3)  weakly-consolidated,  (4)  non-conformable,  (5) 
broken. 

With  case  (1)  or  a  homogeneous  roof  stratum,  either  massive  or  in  a 
sufficiently  thick  bed  to  act  as  such,  the  lines  of  vertical  pressure  far 
3  33 


34  MINING    WITHOUT    TIMBER 

above  ab  tend  to  combine  themselves  into  resultants  which  follow  a 
surface  acb  and  throw  the  downward  pressure  onto  the  walls  at  a  and  6. 
The  resultant  surface  takes  the  form  of  an  arch,  over  a  tunnel,  or  of  an 
arch  with  domed  ends,  in  a  room  of  limited  length.  This  means  that  the 
sub-arch  block  acb  is  all  the  weight  that  has  to  be  supported  to  maintain 
the  roof  intact,  and  that  its  stability  depends  first,  on  its  strength  as  a 
beam  of  continuous  width  to  bear  its  own  weight  across  the  span  ab  and 
second,  on  its  being  held  in  place  by  the  tensile  strength  of  the  rock  area 
along  the  arch,  or  potential  surface  of  fracture,  acb.  In  case  (1)  the 
natural  arching  is  usually  sufficiently  convex  so  that  the  sub-arch-block 
has  sufficient  depth  cf  to  make  is  self-sustaining  as  a  beam  across  ab 
except  in  soft  rocks  like  certain  shales  which  may  not  only  need  the  sup- 
port of  a  cap  like  ab  but  also  must  be  lagged. 

In  old  mine  workings  where  the  sub-arch  block  acb  has  fallen  out  so 
that  the  shape  of  the  natural  surface  of  equilibrium  acb  can  be  discerned, 
it  appears  as  an  arch  whose  proportions  vary  with  the  width  of  the  room 
and  the  nature  of  the  roof.  Fayal  gives  as  working  rules  for  limited 
areas  like  rooms: 

If  w  =  width  of  room  (as  ab  in  Fig.  13) ; 
h  =  height  of  arch  (as  c/in  Fig.  13); 

If  w  is  less  than  6  ft.,  h  may  be  as  much  as  2w  (Fayol's  first  rule) ; 

If  w  is  more  than  6  ft.,  h  may  be  as  much  as  4w  (Fayol's  second  rule). 

In  railroad  or  mine  tunnels,  a  homogeneous  roof  can  be  made  self-sus- 
taining by  excavating  it,  at  the  start,  along  the  natural  arch  form.  In 
the  rooms  of  coal  seams,  however,  or  in  iron-ore  beds,  the  sub-arch  block 
must  be  sustained  intact  until  the  mineral  beneath  is  removed  and  the 
room  abandoned.  The  tensile  strength  of  the  arched  surface  is  seldom 
sufficient  to  accomplish  this  unaided,  except  in  narrow  rooms.  In  wider 
rooms,  a  cross-beam  ab,  or  one  or  more  props  like,//1'  must  be  put  in  whose 
strength,  however,  need  only  equal  the  difference  between  the  weight  of 
the  sub-arch  block  acb  and  the  tensile  strength  of  the  arched  surface  acb, 
provided  that  ff  is  inserted  before  the  surface  acb  has  begun  to  fracture. 
Should  the  latter  accident  have  taken  place,  the  weight  of  the  whole  sub- 
arch  block  may  have  to  be  sustained  by  props  and  thus  a  heavy  unneces- 
sary expense  be  incurred. 

With  case  (2)  or  where  the  roof  is  in  beds  thinner  than  the  sub-arch 
block  so  that  bedding  planes  like  hk  and  mn  (Fig.  13)  intersect  the  surface 
acb,  a  different  condition  arises  from  case  (1).  It  is  evident  that  now 
the  sub-arch  block  instead  of  being  a  single  stone  beam  acb  is  divided 
into  three  stone  beams  ahkb,  h'mnk'  and  m'dd'n' ',  so  that  for  a  self-sustain- 
ing roof,  the  lowest  beam  ahkb  must  be  strong  enough  to  sustain  the 
weight  of  the  two  beams  above  it,  the  central  beam  h'mnk'  must  sustain 
the  top  beam  m'dd'n'  and  the  tensile  strength  of  the  sub-arch  surface  acb 
must  be,  as  before,  sufficiently  strong  to  hold  UD  the  whole  sub-arch 


PRINCIPLES    FOR    CONTROLLING    EXCAVATIONS  35 

block.  It  is,  therefore,  likely  that  a  room  would  need  stronger  props  in 
case  (2)  than  in  case  (1)  because  the  lowest  sustaining  beam  of  case  (2) 
has  a  depth  at  the  middle  of  ha,  which  is  only  a  fraction  of  the  correspond- 
ing depth  cf  for  case  (1),  and  the  cross  breaking  strength  of  a  beam 
increases  directly  as  the  square  of  the  depth.  Also  we  now  do  not  have 
a  uniform  tensile  strength  for  one  surface  of  fracture  acb,  but  a  different 
strength  for  each  of  the  three  beds  which  acb  intersects.  Hence  for  case 
(2)  we  have  to  ascertain  both  the  cross  breaking  and  the  tensile  strengths 
of  all  beds  in  the  sub-arch  block  before  we  can  ascertain  how  much  prop- 
ping is  required  to  sustain  the  roof  across  a  room  of  a  given  width.  A 
roof  of  an  elastic  nature  like  slate  may  at  first  simply  bow  downward 
from  an  excess  of  pressure  instead  of  fracturing  as  a  beam.  This  may 
cause  it  to  fail  by  shear  at  the  abutments.  For  the  maximum  strength 
of  a  roof  it  is  important  to  exclude  water  from  the  bedding  planes  in  order 
to  prevent  the  slipping  and  weakness  caused  by  its  presence. 

Case  (3)  often  occurs  in  coal  mines  where  the  roof  stratum  is  "  clod  " 
or  a  kind  of  soft  shale  containing  concretions  of  considerable  size.  A 
common  device  is  to  leave  the  upper  layer  of  the  coal  seam  under  it  which 
then  acts  as  the  lowest  beam  ahkb  of  case  (2)  to  partially  sustain  the 
elod-stratum.  Where  all  the  seam  must  be  removed  beneath  the  clod, 
the  roof  can  only  be  kept  intact  by  excavating  the  mineral  with  little  or 
no  blasting  and  keeping  the  supports  close  to  the  working  face.  Props, 
cross-pieces  and  lagging  may  all  have  to  be  used.  If  the  clod  stratum  is 
thicker  than  the  room's  natural  arch,  masses  may  fall  out  from  above 
the  surface  acb,  after  the  sub-arch  block  has  been  taken  down,  so  that 
roofs  must  be  arched  higher  than  the  clod  in  order  to  stand  permanently 
unsupported. 

Where  the  roof  is  a  weakly  consolidated  stratum  of  more  uniformly 
sized  stones,  like  a  conglomerate,  the  problem  of  support  is  similar. 
Practically  the  whole  weight  of  the  sub-arch  block  must  be  held  in  place 
by  artificial  supports,  and  in  addition  the  beam  acb  itself  must  be  rein- 
forced by  cross-beams  of  props  or  by  both.  Where  the  roof  stratum  is 
so  weakly  consolidated  as  to  be  incoherent  it  requires  close  lagging,  and 
where  quite  loose  an  advance  can  only  be  made  by  driving  fore-poles 
ahead  of  the  timber  sets  right  up  to  the  working  face.  Loose  sand,  if 
dry  or  only  moist,  can  be  sustained,  like  loose  gravel,  by  close  fore-poling, 
but  if  it  is  wet  enough  to  flow  freely  like  quicksand  the  case  is  hopeless 
except  by  the  use  of  some  such  system  as  that  of  the  pneumatic  shield 
recently  employed  in  the  Hudson  river  tunnels  at  New  York. 

It  is  evident,  however,  that  while  the  quicksand  roof  of  a  railroad 
tunnel  might  be  penetrated  and  sustained  by  the  expensive  pneumatic 
shield  and  its  follower,  a  cast-iron  tube  lining,  such  a  device  would  be 
commercially  unpractical  for  ordinary  ore  deposits.  For  the  latter  the 
only  hope  for  overcoming  quicksand  is  sufficient  drainage  so  that  the 


36 


MINING    WITHOUT    TIMBER. 


sand  loses  its  fluidity  and  takes  the  compact  condition  of  its  merely  moist 
state.  If  drainage  of  the  quicksand  covering  is  not  feasible  and  the  ore 
body  cannot  be  mined  by  some  subaqueous  method,  it  is  worthless,  at 
was  recently  proved  for  a  huge  hematite  deposit  under  a  swamp  on  the 
Mesabi  range,  Minn.,  which  was  abandoned  after  wasting  a  large  sum  in 
attempting  to  open  a  mine  in  it. 

Even  if  the  bed  or  pocket  of  quicksand  does  not  rest  directly  on  the 
ore  body,  but  is  separated  from  it  by  a  rock  stratum,  great  care  has  to  be 
taken  against  it.  The  only  safe  plan  is  to  open  the  mine  excavations 
of  small  size  and  with  sufficient  support  to  keep  the  rock  roof  intact,  for 
otherwise  vertical  cracks  may  develop  reaching  to  the  quicksand. 
When  the  quicksand  once  begins  to  flow  into  the  mine,  the  results  may 


Fia.  14. — Non-conformable  roof. 

be  far-reaching  for  its  escape  from  its  matrix  may  mean  the  collapse  of 
the  latter  and  consequent  disastrous  movements  of  the  whole  overlying 
cover. 

In  case  (4)  the  non-conformable  roof  strata  may  dip  in  any  direction 
with  reference  to  the  underlying  mineral  seam.  If  the  roof  strata  strike 
along  the  long  axis  of  the  room  as  in  the  cross  section  of  Fig.  14,  then  it  is 
evident  that  conditions  will  not  produce  an  arch  of  fracture  as  in  Fig.  13. 
The  upper  stratum  g'gpv  is  entirely  above  the  room  opening  and  bridges 
it  slantingly  from  one  side  to  the  other,  while  the  two  lower  strata,  vpv'q' 
and  q'v'q,  have  their  lower  ends  unsupported  and  projecting  like  canti- 
lever beams.  Then  the  natural  surfaces  of  fracture  will  be  normal  to  the 
bedding  planes  and  will  be  qq'  for  the  lowest  and  q'v  for  the  middle  stratum. 
The  tensile  strength  of  surface  qq'  must  be  enough  to  hold  the  weight  of 
projection  q'v'q  and  any  unbalanced  pressure  from  above,  while  surface 
vq'  must  hold  its  end  vpv'q'  and  any  weight  above.  A  line  of  props  at  t 
strong  enough  to  sustain  the  excess  of  strain  over  the  resisting  strength 
of  surface  qq',  will  hold  up  the  roof  without  a  second  line  at  t  provided 
that  surface  v'p'  is  strong  enough  to  sustain  the  weight  on  it  from  the 
projection  p'pv'. 


PRINCIPLES    FOR    CONTROLLING    EXCAVATIONS  37 

As  the  dip  of  the  roof  beds  increases,  the  strain  on  the  surface  of 
fracture  qq'v  becomes  more  tensile  than  cross-breaking  until  with  vertical 
beds  the  strain  is  all  tensile.  In  the  last  case,  the  weight  to  be  sustained 
by  each  stratum  is  its  block  below  a  natural  arch  of  fracture  across  the 
room,  which  is  differently  proportioned  for  vertical  beds  than  is  acb  of 
Fig.  13  for  horizontal  beds. 

Where  the  strike  of  the  inclined  beds  of  the  roof  is  across  instead  of 
along  the  room  beneath,  we  have  a  mixture  of  the  cases  illustrated  by 
Figs.  13  and  14.  Each  bed  can  first  be  considered  separately  as  forming 
a  single  sub-arch  beam  whose  side  elevation  is  acb  in  Fig.  13.  Each  bed 
must  then  be  calculated  separately  both  for  the  self-sustaining  power  of 
its  sub-arch  beam,  across  the  span  of  the  room,  and  for  that  of  its  tensile 
surface  acb.  A  bed  may  then  be  artificially  supported  if  necessary,  by 
prop  ff  or  cap  ab.  If  Fig.  14  be  assumed,  for  this  case  only,  to  be  the 
longitudinal  section  of  the  room  whose  cross-section  is  Fig.  13,  we  see  that 
a  cross  bed  like  -vp-v'q'm&y  have  the  same  breaking-off  action  on  a  lower 
bed  q'v'q  as  has  just  been  discussed  in  the  last  paragraph,  and  supports 
must  be  modified  accordingly. 

The  broken  roof  of  case  (5)  may  arise  from  planes  of  faulting,  fractur- 
ing, jointing,  etc.  If  the  breaking  planes  are  parallel  or  in  one  general 


FIG.   15. — Roof-over  inclined  room. 

direction,  we  can  handle  the  roof  as  suggested  for  case  (4) .  If  the  planes 
are  in  several  directions  so  as  to  cut  the  roof  into  monoliths,  the  support 
of  each  block  will  have  to  be  studied  separately.  Where  a  roof  monolith 
is  of  indefinite  height,  we  may  illustrate  it  by  Fig.  13  with  ab  its  length 
and  acb  the  section  of  its  natural  surface  of  fracture,  which  will  be  of 
dome  shape,  so  that  only  the  support  of  the  sub-arch  portion  acb  has  then 
to  be  considered.  When,  however,  the  roof  monoliths  are  broken  also 
by  a  plane  in  a  horizontal  direction,  like  mn  in  Fig.  13,  so  as  to  become 


38  MINING    WITHOUT   TIMBER 

free  blocks  like  amnb,  they  can  only  be  kept  in  the  roof  by  sustaining 
their  entire  weight  artificially,  and  fore-poling  will  have  to  be  used  for 
excavating  beneath  a  roof  surface  containing  them. 

(6)  Roof  Over  an  Inclined  Room. — This  case  occurs  in  mining  seams  on 
a  dip  which  may  vary  up  to  90  deg.  from  the  horizontal.  Let  abb'a',  Fig.  15, 
represent  the  cross-section  of  a  room  in  a  seam  of  the  thickness  bb',  which 
has  the  usual  horizontal  floor  aa'  for  tramming.  It  is  evident  that  the 
principles  of  roof  support  similar  to  the  previous  case  of  horizontal  rooms 
apply  here,  but  the  action  of  the  superincumbent  weight  in  the  roof  is 
affected  by  the  angle  of  dip.  Thus  in  the  diagram  of  Fig.  15,  if  W  =  super- 
incumbent weight;  6  =  angle  of  dip;  N  =  normal  pressure  on  roof; 
T  =  tangential  pressure  on  roof;  then 

N=  W  cos  0 (1) 

and  T=  W  sin  0 (2) 

For  homogeneous  strata  the  weight  of  the  overlying  formation  would 
be  thrown  onto  the  pillars  at  a  and  b  and  the  potential  surface  of  fracture 
would  be  the  arch  acb.  Thus  the  span  ab  has  only  to  sustain  the  normal 
pressure  of  the  sub-arch  block  acb  acting  both  in  tension  on  the  surface 
acb  and  in  cross-breaking  strain  on  the  beam  acb  as  described  for  case  (a) 
in  Fig.  13.  The  back  of  the  ore  should  also  fall  on  the  arch  line  bdb' 
instead  of  a  straight  line  from  b  to  &'.  A  prop  to  hold  up  the  roof  will 
be  subjected  to  the  least  pressure  and  be  of  shortest  length  if  it  is  placed 
in  a  line  gfe  drawn  normal  to  the  hanging  wall  from  the  center  of  gravity 
of  the  sub-arch  block  at  g.  Because  of  possible  shrinkage  of  prop  or 
movement  of  ground,  however,  which  would  cause  a  normal  prop  to  fall 
out,  the  usual  practice  is  to  incline  it  about  10  deg.  downward  from  the 
normal  line  as  ff.  The  sub-arch  block  acb  can  also  be  sustained  by  a 
cap  ab  from  the  back  to  the  floor,  or  by  both  prop  and  cap.  With  the 
roof-strata  bedded  parallel  to  the  seam,  the  surface  of  fracture  assumes 
the  stepped-arch  form  ahh'mm'k'kb.  In  comparing  the  strains  and  the 
support  of  the  bedded  roof,  as  well  as  of  the  weakly-consolidated,  of  the 
non-conformable  and  the  broken  roofs  with  those  of  the  homogeneous 
roof,  the  same  differences  arise  as  already  explained  for  case  (a) 

CONTROL  OF  THE  OVERLYING  FORMATIONS 

It  is  evident  that  when  part  of  a  bed  is  removed,  the  balance  left  as 
pillars  must  sustain  the  whole  overlying  formation.  There  are  three 
factors  that  enter  into  pillar  calculations,  the  roof,  the  pillars  or  sides 
and  the  floor.  The  stability  of  the  room  does  not  depend  alone  on  the 
strength  of  the  pillars  as  columns  for  an  excess  of  pressure  may  force  a 
sound  pillar  into  a  roof  or  floor  of  insufficient  compressive  strength  and 
cause  a  settling.  This  happens  with  materials  like  clay  which  are  hard 


PRINCIPLES    FOR    CONTROLLING    EXCAVATIONS  39 

when  dry  and  become  soft  when  moist,  so  unless  they  can  be  kept  dry 
during  mining,  the  pillar  calculations  must  guard  against  their  moist 
state.  An  excess  of  pressure  on  a  plastic  floor  will  cause  it  to  spread 
laterally  and  rise  from  under  the  lower  periphery  of  the  pillar,  thus 
exerting  a  horizontal  rending  force  on  the  latter  which  tends  to  disrupt 
its  edges. 

Any  downward  bowing  of  an  elastic  roof  over  the  rooms  must  be 
compensated  for  by  an  upward  bowing  over  the  interior  of  the  pillars. 
This  causes  an  oblique  pressure  at  the  upper  edges  of  the  latter  which 
tends  to  shear  them  off  as  the  roof  bends  more  and  more.  The  obliquity 
of  this  roof  pressure  on  the  pillar  edges  is  also  often  increased  by  a  rolling 
floor. 

Thus  a  mine  floor  and  roof  act  not  only  vertically  on  the  areas  of 
contact  with  the  pillars,  but  also  laterally,  while  the  bowing  of  the  roof 
produces  strains  parallel  to  the  strata  that  tend  to  separate  them  along 
their  bedding  planes,  and  thus  weaken  the  cross  breaking  strength  of  the 
roof.  For  these  reasons,  a  mine  pillar  will  stand  best  and  can  be  made  of 
the  minimum  volume  when  its  .base  and  capital  meet  floor  and  roof  in 
broadly  spreading  tangential  curves,  which  are  concave  in  profile. 

Sometimes  the  roof  and  floor  beds,  in  direct  contact  with  the  seam, 
are  themselves  quite  hard,  but  so  thin  that  they  bend  and  transmit  the 
pressure  on  the  pillars  to  an  adjoining  soft  stratum  and  force  it  out 
through  any  fissures  that  may  be  in  the  roof  or  floor.  When  the  pillars 
themselves  are  too  weak  for  the  pressure,  what  is  called  a  "squeeze" 
(failure  of  pillars)  begins,  by  a  shelling  off  of  the  outer  surface,  and  later 
a  collapse  occurs,  which  may  be  a  gradual  sinking,  with  elastic  strata 
like  some  coals  and  shales  or  a  sudden  fracturing  in  masses  with  hard 
blocky  rock-like  limestone  or  quartz.  Some  substances,  like  coal,  pyrite 
and  easily  weathered  rocks,  loose  strength  on  exposure  to  mine  air  and 
this  fact  must  be  considered  if  durable  pillars  are  to  be  made  of  them. 
The  minimum  fraction  of  a  bed  necessary  to  leave  for  pillars  may  be 
thus  calculated: 

Let  x    =  fraction  of  area  to  be  left  in  pillars; 
h   =  depth  of  cover  in  feet; 

w   =  specific  weight  of  cover  in  pounds  per  cu.  ft.; 
s   =  ultimate  compressive  strength  in  pounds  per  square  foot  of 

least  resistant  stratum  adjoining  pillar; 
m  =  factor  of  safety, 

Then,  weight  held  by  1  sq.  ft.  of  seam  =  hw,  and  compressive  strength 
of  corresponding  pillar  =  xms,  hence  hw  =  xms. 

hw 

or  x  =  — (3) 

ms 

If  excavation  is  inclined  at  an  angle  as  in  Fig.  15,  then  the  pressure  is 


40  MINING    WITHOUT    TIMBER 

hw  cos  0 

hwcosO,  so  that  hw  cos  0  =  xms  or  x  =  -        — (4) 

ms 

It  is  difficult  to  get  the  real  strength  of  the  floor,  pillar  and  roof  beds 
because  the  beds  themselves  are  seldom  free  from  planes  of  weakness 
which  would  not  be  appreciable  in  the  small  blocks  that  must  be  used  in 
the  testing  machines  for  compression,  tension  or  shear.  For  this  reason 
the  factor  of  safety  ra  of  equations  (3)  and  (4)  is  taken  at  from  2  to  10, 
varying  with  the  nature  of  the  strata  and  of  the  mine  layout. 

It  is  only  by  close  watching  on  the  changing  conditions  that  move- 
ments of  the  formation  over  wide  excavations  can  be  prevented  even  in 
well  laid  out  mines.  An  incipient  "  squeeze  "  of  pillars  may  sometimes 
be  checked  by  building  up  stone-filled  wooden  cribs  along  their  edges, 
but  this  remedy  may  merely  shift  the  pressure  and  transfer  the  "  squeeze  " 
elsewhere.  Often  it  is  better  to  localize  rather  than  to  attempt  to  support 
a  squeeze  and  this  can  be  affected  by  allowing  the  roof  to  cave  over  the 
disturbed  section,  assisting  the  fall  where  necessary  by  blasting  the  roof 
and  pillars.  The  volume  of  roof  thus  made  to  fall  will  be  that  under  the 
dome  of  fracture  as  acb  of  Fig.  13,  the  span  ab  in  this  case  not  being  the 
width  of  a  single  room,  but  of  the  whole  disturbed  section.  If  the  seam 
is  thin  in  proportion  to  the  height  of  the  falling  dome,  the  broken  rock, 
as  it  occupies  more  space  than  when  solid,  will  fill  up  the  space  under  the 
surface  of  fracture  and  form  a  sufficient  support  to  prevent  any  further 
strain  on  the  overlying  formation. 

The  caving  of  the  roof  over  the  disturbed  area  is  also  a  remedy  for 
"creep"  (oozing  of  roof  or  floor  into  excavations),  but  if  the  ground 
surface  is  to  remain  intact  a  safer  plan  is  to  fill  the  excavation  solid  with 
rock.  Where  a  supply  of  fine  material  like  mill  tailing  or  sand  can  be 
obtained  cheaply,  the  filling  is  best  done  by  mixing  it  with  water  and 
running  it  into  the  workings  through  pipes  by  the  flushing  system  of  the 
Pennsylvania  anthracite  regions  as  described  in  Examples  59  and  60. 

The  caving  of  the  roof,  locally,  by  blasting  can  be  easiest  affected  by 
reversing  the  methods  already  explained  for  roof  support.  If  pulling'  or 
blasting  out  all  artificial  supports  does  not  bring  down  the  roof,  any  rock 
pillars  in  the  area  should  be  drilled  and  blasted  by  simultaneous  firing. 
The  next  procedure  is  to  drill  holes  into  the  roof  so  as  to  cut  a  groove 
around  the  springing  line  of  the  dome  acb  in  Fig.  13.  The  work  of  the  drill 
men  around  the  edges  of  the  excavation  will  be  safe  and  the  circumferen- 
tial groove  can  thus  easily  be  widened  and  carried  higher  until  the  central 
bell  of  the  sub-arch  dome  has  so  much  of  its  sustaining  surface  acb  cut 
away  that  it  drops  out. 

EFFECT  OF  CAVING  ON  OVERLYING  OBJECTS 

In  working  superimposed  beds  simultaneously,  it  is  necessary  to 
determine  the  proper  relative  position  of  pillars  in  the  various  beds. 


PRINCIPLES    FOR    CONTROLLING    EXCAVATIONS 


41 


Pillars  must  also  be  located,  in  caving  mines,  where  it  is  desired  to  pro- 
tect valuable  surface  structures.  In  modern  coal  mining,  both  the  long- 
wall  and  usually  the  room  and  pillar  method  involve  the  caving  of  the 
excavations. 

How  far  up  an  underground  subsidence  will  reach  depends  on  a  num- 
ber of  conditions,  such  as  area,  height  and  manner  of  making  of  excava- 
tion, nature  of  overlying  formation,  presence  of  faults  and  dikes,  etc.  By 
Fayol's  second  rule,  the  height  affected  by  subsidence  would  not  exceed 
four  times  the  width  of  the  excavation,  but  this  only  holds  good  for  a 
limited  area  whose  sub-arch  roof  block  can  scale  off  at  leisure.  When 
large  areas  are  excavated,  complex  stresses  arise  which  are  apt  to  cause 
sudden  irresistible  strains  on  the  roof  which  cause  it  to  develop  long  cracks 
and  fractures  analogous  to  faults.  If  the  overlying  strata  contain  many 
strong  rock  beds,  these  may  act  as  beams  which  rest  on  the  broken  caved 
formation  beneath  them  and  prevent  any  effect  above.  Thus  at  Sunder- 
land,  England,  where  half  of  the  strata  are  hard  rock,  coal  seams  have 


FIG.  16. — Effect  of  excavation  on  overlying  bed  and  on  surface. 

been  mined  and  caved  at  the  depths  of  1600  ft.  without  affecting  the 
surface.  In  the  Transvaal  gold  beds,  dipping  at  around  40  deg.,  caves  may 
occur  over  areas  of  several  acres  at  depths  over  1000  ft.  without  surface 
movement.  With  a  formation  of  soft  friable  strata,  like  shale  .or  glacial 
drift,  however,  there  is  nothing  to  arrest  a  subsidence  beneath,  and  under 
such  roofs  the  effect  of  caving  coal  mines,  2000  ft.  deep,  has  depressed 
surface  structures. 

Fayal's  third  rule  applies  to  excavations  of  large  area  and  is  "where 
the  area  is  infinite  and  the  beds  are  chiefly  sandstone  with  a  dip  less  than 


42  MINING    WITHOUT    TIMBER 

40  deg.,  the  height  of  the  zone  of  subsidence  is  less  than  200  times  the 
height  of  the  excavation."  This  means  that  the  caving  of  an  excavation, 
6  ft.  high,  would  not  affect  the  surface  if  over  1200  ft.  below  it.  The  third 
rule  is  based  on  the  height  of  excavation  rather  than  on  its  width,  like  the 
other  rules,  and  depends  on  the  principle  already  mentioned  that  the 
strong  strata  tend  to  rest  solidly,  ultimately,  on  the  caved  ground  below. 
Subsidence  does  not  break  strata  perpendicular  to  their  bedding 
planes.  For  denning  the  disturbed  area  over  excavations  under  unbroken 
stratified  formations  two  rules  are  used,  the  first  for  slightly  and  the 
second  for  steeply  dipping  roofs.  Thus  in  Fig.  16, 
if  D  =  dip  of  roof  strata  in  degrees 

A    =  dip  of  angle  of  fracture, 
for  roofs  under  30  deg.  dip  Richardson1  gives, 

A  =90  deg.- 1/2  D (5) 

which  signifies  that  the  plane  of  fracture  e  f  (Fig.  16)  of  bed  ab  lies  half 
way  between  the  vertical  and  the  plane  eg  (normal  to  the  dip  line  of  the 
roof). 

For  roofs  over  30  deg.  dip  Hausse1  gives, 

tan  A  =  cotan  2D  +3  cosec  2D (6) 

Formula  (6)  gives  for  a  30-deg.  roof  only  a  slightly  larger  angle  of  fracture 
than  formula  (5),  but  as  the  dip  gets  steeper  the  difference  between  the 
two  formulas  steadily  increases  while  a  maximum  A  is  reached  with 
formula  (6)  when  D  is  between  50  deg.  and  60  deg.  as  shown  in  the 
following  table: 

Angle  A,  degs. 

Dip  D  Formula  (5)  Formula  (6) 

Odeg.  90  90.0 

10  deg.  85  85.2 

20  deg.  80  80.5 

30  deg.  75  76.2 

40  deg.  70  73.0 

50  deg.  65  70.8 

60  deg.  60  71.0 

70  deg.  55  74.0 

80  deg.  50  80.8 

90  deg.  45  90.0 

These  formulae  can  only  be  considered  as  general  guides  to  the  prob- 
able location  of  the  plane  of  fracture  and  they  must  be  modified  in  practice 
by  a  consideration  of  the  surface  topography,  of  the  structure  of  the 
formation  and  of  natural  breaks  like  joints  and  faults.  Where  thick 
dikes  cut  across  the  roof  strata,  the  plane  of  fracture  is  more  apt  to  follow 
along  the  surface  of  the  dike  than  to  break  it. 

^ng.  and  Min.  Journal,  Aug.  3,  1907,  p.  196. 


PRINCIPLES    FOR    CONTROLLING    EXCAVATIONS  43 

Protection  of  Surface. — The  practical  use  of  these  formulae  is  shown 
in  Fig.  16  where  it  is  desired  to  protect  the  building  at  /&'  when  mining 
the  veins  ab  and  cd.  Here  h'f  and  hf  are  planes  drawn  parallel  to  the 
plane  of  fracture  ef  and  their  intersection  with  the  beds  defines  the  inside 
boundaries  of  the  pillars  e'e  and  h'h.  The  margin  of  safety  to  be  left 
around  these  inside  limits  of  the  pillars  for  "draw"  varies  with  the 
importance  of  the  building  and  how  closely  the  strata  have  been  observed 
to  follow  the  fracturing  formulae. 

Protection  of  Overlying  Beds. — Where  the  veins  cd  and  ab  are  being 
mined  simultaneously,  it  is  evident  that  the  pillars  to  be  left  in  cd  to 
protect  pillar  e'e  must  not  be  the  ground  vertically  beneath,  as  hk,  but 
that  enclosed  between  the  same  planes  of  roof  fracture  as  h'h  with  a  due 
allowance  added  for  "draw."  In  excavating,  also,  the  direction  of  the 
roof  fracture  ef  must  be  taken  instead  of  the  vertical  plane  as  the  guide 
to  relative  operations  in  the  upper  and  the  lower  beds.  Thus  for  safety 
the  bed  ab  would  be  stopped  "  ahead  "  (measuring  from  the  plane  ef)  of 
bed  cd;  except  in  the  case  where  cd  was  being  filled,  when  the  slight  sub- 
sidence of  the  floor  of  ab,  caused  by  the  settling  of  cd  (when  "  ahead  ")  on 
its  filling,  would  render  the  breaking  of  ab  easier. 

In  mining  the  superincumbent  parallel  anthracite  seams  of  the  Lehigh 
Valley  Coal  Co.,  by  the  room  and  pillar  system,  the  pillars  must  overlay 
each  other  when  the  parting  is  thin.  A  neglect  of  this  precaution,  with 
the  usual  parting,  is  liable  to  result  in  the  squeezing  of  the  overlying  pillars 
down  into  the  rooms  of  the  seam  below.  When  the  parting  is  over  40  ft. 
thick,  however,  it  is  only  necessary  to  have  the  panel  pillars  (at  ten-room 
intervals)  of  adjoining  seams  superincumbent,  and  to  lay  out  the  entries 
and  room  axes  of  both  seams  approximately  parallel  to  each  other;  in 
this  way  the  work  in  different  seams  can  be  pursued  more  independently 
and  just  as  safely. 

Shaft  Pillars. — The  same  principles  and  formulae  can  be  applied  to 
the  design  of  pillars  for  protection  of  shafts.  In  Fig.  16  the  vertical  shaft 
fd  will  need  a  pillar  in  each  seam  extending  to  the  intersection  with  the 
plane  of  fracture  passing  through  the  shaft  collar  at  /.  Thus  the  mini- 
mum upper  limit  of  these  pillars  must  be  at  e  and  h,  which  for  considerable 
depth,  would  mean  many  hundred  feet  away  from  the  shaft.  But  this 
involves  only  a  moderate  loss  of  ore  because  the  pillar  may  be  narrow 
and  need  extend  only  a  short  distance  down  the  dip  to  6  and  d.  The 
distances  b'b,  d'd  and  the  width  of  the  shaft  pillar  along  the  strike  of  the 
seam  may  be  estimated  by  formula  (4) . 

For  inclined  shafts  following  the  mineral  seam,  the  protecting  pillars 
should  be  continuous  strips  on  each  side  with  break-throughs  only  for  the 
loading  stations.  The  width  of  these  strips,  if  estimated  by  formula  (4), 
should  increase  gradually  from  the  surface  downward.  Although  this 


44 


MINING    WITHOUT   TIMTJER 


last  requirement  is  seldom  fulfilled  in  practice,  it  gives  the  minimum  loss 
of  ore  for  the  maintenance  of  a  stable  roof. 

SUPPORT  OF  EXCAVATIONS 

Curved  Sections. — A  tunnel  section  may  be  supported  by  the  circular 
lining  1  (Fig.  17)  against  external  pressure  from  any  direction  since  the 
portion  of  the  ring  taking  the  ground  pressure  will  be  an  arch  to  transmit 
its  load  to  the  balance  of  the  ring  acting  like  arch  abutments.  If  we 
consider  only  the  keeping  open  of  a  given  area  with  the  least  material,  the 
circular  lining  may  be  replaced  with  advantage  by  the  elliptical,  when  the 
pressure  is  greater  in  one  direction  than  in  another,  by  placing  the  long 
axis  of  the  ellipse  parallel  to  the  direction  of  greatest  pressure.  Thus  if 
the  greatest  pressure  comes  from  roof  or  floor,  the  ellipse  should  be  verti- 
cal as  2  in  Fig.  17,  and  if  from  the  sides,  the  ellipse  should  be  horizontal  as 


FIQ.  17. — Tunnel  sections. 

in  3.  The  circular  lining  is  most  economical  when  the  external  pressure 
is  equally  distributed,  or  where  it  comes  in  an  oblique  direction,  for  an 
oblique  ellipse  would  be  generally  unsuitable  for  use.  The  oblique  pres- 
sure is  apt  to  occur  when  driving  along  the  strike  of  inclined  beds.  Other 
considerations,  besides  economy  of  lining,  usually  prevail  in  practice  so 
that  circular  sections  are  less  used  than  elliptical  ones,  which  fit  cars 
more  closely  in  transportation  tunnels,  or  egg-shaped  ones,  like  4,  which 
have  a  lesser  hydraulic  gradient  for  water  conduits  or  drains.  The 
material  most  used  for  curved  linings  is  cast  iron  or  some  kind  of  masonry, 
though  steel  shapes  are  also  formed  to  fit,  and  timber  polygons  to  approxi- 
mate curved  sections.  To  merely  support  the  ground,  it  is  clear  that 
only  that  part  of  the  tunnel  section  need  be  lined  which  has  weak  walls 
so  that  we  see  in  practice  linings  on  the  roof  alone,  on  the  roof  and  one 


PRINCIPLES    FOR    CONTROLLING    EXCAVATIONS 


45 


side,  or  on  three  sides,  the  ends  of  the  lining  resting  in  each  case  against 
abutments  of  solid  rock. 

Rectangular  Sections. — The  greatest  available  area  in  transportation 
tunnels  for  the  minimum  volume  of  excavation  is  obtained  by  using  the 
rectangular  instead  of  the  curved  section.  Ordinary  brick  or  stone 
masonry,  having  little  tensile  strength,  is  unsuitable  for  lining  any  part 
of  the  rectangular  section  subject  to  cross-breaking  strains.  Therefore 
it  is  not  used  except  for  side  walls.  Timber  or  steel  beams  and  re-en- 
forced concrete  are  the  common  linings  for  rectangular  sections.  With 
a  weak  roof  and  strong  sides  only  the  piece  ab  (Fig.  18),  which  is  called  a 
cap  or  a  "quarter-set"  is  put  in;  with  both  roof  and  one  side  weak  the 
cap  ab  and  the  post  be,  called  a  "  half  set, "  are  needed;  with  roof  and  both 


FIG.  18. — Tunnel  lining  with  timber. 

sides  weak  a  cap  and  two  side  posts,  or  a  "three-quarter  set,"  is  used; 
while  with  four  weak  walls  a  cap,  two  posts  and  a  floor  sill,  or  a  "full  set," 
is  required. 

The  attempt  will  not  be  made  here  to  discuss  methods  of  framing  except 
as  they  are  affected  by  ground  pressure.  With  a  predominating  vertical 
pressure  a  good  joint  for  square  timber  is  at  a  (Fig.  18),  and  for  a  round 
timber  at  6.  Where  the  main  pressure  is  horizontal  a  joint  for  square 
timber  is  shown  at  d  and  one  for  round  timber  at  c.  A  rectangular  frame 
can  resist  pressure  acting  parallel  to  its  sides,  but  tends  to  collapse  under 
oblique  pressure.  It  is  to  make  them  more  stable  under  oblique  pressure 
that  tunnel  sets  have  outward-battered  instead  of  vertical  posts. 

Slope  Sections. — In  keeping  open  the  large  stopes  of  some  metal  mines 
with  the  framed  cubical  cells  of  the  square-set  system,  the  same  precau- 
tions of  properly  designed  joints  and  of  uniformly  spaced  points  of  contact 
with  the  surrounding  rock  must  be  observed.  Miners  have  found  to  their 
sorrow  that  it  is  useless  to  attempt  to  keep  open  square-setted  stopes 
under  oblique  pressure  unless  diagonal  braces  (like  bd)  are  inserted  par- 


46  MINING    WITHOUT    TIMBER 

allel  to  the  direction  of  pressure,  for  transforming  the  unstable  squares 
of  the  frame  into  stable  triangles. 

ZONES  OF  FRACTURE  AND  FLOW  AGE 

Wooden  or  steel  frames  will  only  keep  the  peripheral  surface  of 
excavations  intact  against  the  pressure  of  loose  pieces  or  sub-arch  blocks 
like  acb  in  Fig.  13.  For  the  support  of  the  overlying  formation,  even 
masonry  is  only  of  limited  commercial  utility,  therefore  rock  pillars  or 
filling  with  waste  must  be  relied  upon.  Beyond  a  certain  depth,  or  below 
the  "zone  of  fracture"  of  geologists,  we  have  the  "zone  of  flowage," 
where  no  opening  can  be  maintained  permanently  owing  to  the  inability 
of  any  fraction  of  the  rock,  left  as  pillars,  to  sustain  the  superincumbent 
pressure. 

Transposing  formula  (3)  we  have  for  h'  (the  depth  of  the  zone  of 
fracture)  : 

smx 

h  =  ----  , 
w 

but  for  the  zone  of  flowage  both  m  and  x  are  =  1  and  substituting  these 
values  we  have 


(7) 


From  equation  (7)  it  is  evident  that  the  depth  h'  depends  solely  on 
the  compressive  strength  of  the  basal  rock  and  the  specific  gravity  of  the 
overlying  formation.  Assuming  the  specific  weight  w  of  the  earth's  crust 
to  be  150  Ib.  per  cu.  .ft.  and  the  compressive  strength  of  the  basal 
rock  to  be  3,000,000  Ib.  per  sq.  ft.,  we  have  by  substitution  in  (7) 

3,000,000 

h'=-  -  -  —  =  20,000  ft.,  or  about  4  miles. 
150 

Recently  it  has  been  shown  by  Profs.  Adams  and  King1  that  bore- 
holes in  test-cylinders  of  granite  remain  unchanged  even  when  long 
subjected  (at  a  temperature  of  550°  C.)  to  crushing  stresses  of  14,000,000 
Ibs.  per  sq.  ft.  This  indicates  that  the  value  hitherto  assumed  for  s  in 
equation  (7)  is  much  too  small  and  that  the  zone  of  flowage  in  granite 
would  lie  deeper  than  14  miles  from  the  surface. 

*  Journal  of  Geology,  Vol.  XX,  pp.  97-138. 


CHAPTER  V 
PRINCIPLES  OF  MINE  DRAINAGE 

Those  miners  who  talk  much  of  pumping  but  little  of  drainage 
resemble  those  old-fashioned  doctors  who  spend  all  their  time  on  remedies 
and  neglect  diagnosis.  Instead  of  studying  water  conditions  beforehand 
as  a  basis  for  drainage  equipment,  a  too  common  way  is  to  try  to  fit  the 
pumping  plant  to  the  in-flow  after  it  has  appeared.  This  policy  may 
mean  a  drowned  mine,  and  weeks  of  delay  for  the  installation  of  larger 
pumps  and  the  clearing  from  water;  it  may  mean  a  set  of  makeshift 
pumps  of  the  wrong  size  and  of  low  efficiency  which  may  really  be  wholly 
unnecessary  owing  to  the  feasibility  of  natural  drainage. 

Problems  of  drainage  involve  chiefly  the  four  sciences  of  meteorology, 
geology,  hydraulics  and  mechanics.  From  the  first  we  may  determine 
the  quantity  of  rain  likely  to  fall  on  our  mine  watershed;  from  the 
second  the  conditions  affecting  the  behavior  of  underground  water  in 
the  rocks;  from  the  third  the  laws  governing  the  pressure  and  flow  of 
water;  and  from  the  fourth  the  mechanical  methods  of  unwatering. 

ESTIMATE  OF  WATER  TO  BE  DRAINED 

There  are  multitudinous  mineral  deposits,  each  with  a  special  problem 
of  drainage  of  which  only  some  general  features  will  be  discussed  here 
under  three  cases:  I.  Deposits  in  unconsolidated  rock;  II.  Deposits  in 
stratified  rock;  III.  Deposits  in  massive  rock.  For  any  type  the  water 
encountered  in  mining  operations  will  depend  on  four  factors:  (1)  the 
area  of  contributory  watershed;  (2)  the  moisture  falling  on  watershed; 
(3)  the  moisture  percolating  the  surface  of  watershed;  (4)  the  facilities 
for  underground  water  to  enter  the  mine. 

Case  I.  Deposits  in  Unconsolidated  Rocks. — In  Fig.  19  is  shown  a 
cross-section  of  a  gentle  syndinal  rock  trough  ab  filled  with  alluvium  up 
to  the  surface  cd.  It  is  proposed  to  lower  the  "water  table"  or  ground 
water1  level  wt  down  to  sump  s  in  order  to  mine  a  placer  deposit  extending 
from  a  to  b.  The  conditions  which  determine  the  quantity  of  water  to  be 
handled  depend  on  two  items;  namely,  the  quantity  of  ground  water, 
and  its  velocity  of  entrance  into  the  workings.  For  the  first  item  we 
have  to  calculate  the  area,  rainfall  and  percolation  of  the  contributory 
watershed,  while  for  the  second  item  the  fact  that  it  will  be  affected  by 
the  method  of  drainage  will  have  to  be  taken  into  consideration.  The 
area  and  rainfall  are  also  the  basis  of  the  calculations  of  the  water-supply 
i " Ground  Waters"  by  J.  F.  Kemp,  Trans.  Min.  Eng.,  Feb.,  1913. 

47 


48  MINING    WITHOUT    TIMBER 

engineer,  but  the  .latter  reckons  rather  with  the  run-off  than  with  the 
percolation  which  concerns  the  miner. 

With  underground  conditions  as  represented  in  Fig.  19,  the  area 
of  the  contributory  watershed  evidently  extends  in  width  from  e  to  /, 
and  in  length  from  the  sump  s  to  the  head  of  the  valley,  if  cd  is  a  river 
trough,  or  to  the  bounding  contour  of  the  watershed  if  cd  lies  in  a  lake 
basin.  In  general  the  contributory  watershed  is  all  the  ground  area 
that  drains  toward  the  surface  lying  over  the  sump,  wherever  the  surface 
is  connected  with  the  sump  by  pervious  strata  as  in  Fig.  19.  The  depth 


FIG.  19. — Drainage  in  unconsolidated  rock. 

of  current  rainfall  is  recorded  for  most  localities  in  civilized  countries 
at  the  government  meteorological  stations;  and  in  solving  drainage 
problems,  these  records  should  be  scrutinized  for  the  maximum,  mean 
and  minimum  rainfalls  both  by  months  and  years.  From  this  data,  we 
have  the  rainfall  in  the  wettest  year  or  season  in  contrast  with  that  of 
drouths,  but  it  is  important  also  to  note  what  part  of  the  moisture  falls 
as  snow  and  the  melting  seasons  of  the  latter. 

The  whole  rainfall,  however,  does  not  concern  the  miner.  He  is  con- 
cerned only  with  that  fraction  of  it  which  sinks  into  or  percolates  the 
ground  after  evaporation  and  run-off  have  taken  their  tolls.  Then  if 

area  of  a  watershed  =  A  sq.  ft. 

depth  of  moisture  falling  on  a  watershed  —  D  ft. 

volume  of  moisture  falling  on  a  watershed  =  Q  cu.  ft. 

volume  of  moisture  running  off  from  a  watershed  =R  cu.  ft. 

volume  of  moisture  evaporating  from  a  watershed  =E  cu.  ft. 

volume  of  moisture  percolating  a  watershed  =  P  cu.  ft. 

fraction  of  moisture  Q  evaporating,  or  evaporation  factor  —  e 

fraction  of  moisture  Q  running  off,  or  run-off  factor  =  r 

we  have, 

Q  =  E+R  +  P  (1) 

but  E  =  eQ&ndR  =  rQ 

so  substitute  in  (1)  and 

Q  =  eQ+rQ  +  P  or  P  =  Q  (I-e-r)         (2) 
but  Q=AD  (3) 

hence  P  =  AD  (I-e-r)  (4) 

Evaporation  is  dependent  on  the  state  of  the  atmosphere  and  the 
covering  and  texture  of  the  soil.  The  atmosphere  affects  evaporation 


PRINCIPLES    OF    MINE    DRAINAGE  49 

by  its  changes  in  humidity  and  in  movement.  Both  dryness  and  high 
winds  hasten  evaporation  which  is  usually  compared  for  different  atmos- 
pheres by  observing  water  surfaces.  Thus,  in  the  United  States  the 
mean  annual  evaporation  varies  from  40  in.  in  the  Middle  Atlantic  states 
to  50  in.  on  the  Gulf  of  Mexico,  and  95  in.  at  Yuma,  Arizona. 

In  the  same  locality  the  rate  of  evaporation  which  is  approximately 
equal  for  all  bare  soils,  is  greatly  increased  by  a  cover  of  vegetation. 
Thus  a  5-year  trial  at  Geneva,  New  York,  with  an  average  rainfall  of  23.7 
in.,  gave  its  evaporative  factor  (e  in  Equation  (4))  as  0.64  for  bare 
cultivated  soil,  as  0.71  for  bare  undisturbed  soil,  and  as  0.85  for  sod.  Not 
only  the  heat  of  summer  but  its  vegetation  increases  evaporation,  while 
the  ground  surface  in  winter  acts  much  like  bare  soil  unless  covered  by 
snow  or  ice,  the  daily  evaporation  rate  of  which  in  New  England  is 
.02  in.  and  .06  in.  respectively.  A  less  proportion  of  severe  rains  is 
evaporated  than  of  drizzling  rains,  for  as  a  given  area  has  only  a  limited 
rate  of  evaporation  any  excess  moisture  must  either  run  off  or  percolate. 

The  common  method  of  estimating  the  run-off  is  from  measurements 
of  the  quantity  of  water  flowing  in  the  streams  of  the  watershed.  When 
the  bed  of  a  stream  is  once  mapped  in  section,  a  record  of  its  surface- 
height  readings  renders  possible  a  calculation  of  its  sectional  area  which, 
combined  with  corresponding  readings  of  a  current  meter,  gives  the  data 
for  computation  of  flow.  The  percentage  of  rainfall  found  in  streams, 
evaporation  being  neglected,  depends  both  on  the  slope  of  the  surface 
and  on  its  covering.  For  gently  rolling  land  as  in  Iowa,  the  run-off  factor 
(r  in  Equation  (4))  is  0.33,  for  the  rougher  surface  of  the  Middle  Atlantic 
States  it  is  0.40  to  0.50,  while  in  the  mountain  states  of  Colorado  and 
Montana  it  is  0.60  to  0.70.  The  surface  covering  most  favorable 
to  a  heavy  flow  is  frozen  snow  over  which  over  90  per  cent,  of  the  rainfall 
may  run  into  the  streams,  while  the  melting  of  the  winter's  snow  by  warm 
rains  causes  the  freshets  and  floods  of  spring.  Where  the  surface  is 
irregular  so  that  the  rainfall  collects  in  ponds  and  swamps  instead  of 
reaching  streams,  the  run-off  is  lessened,  and  the  evaporation  and  per- 
colation is  correspondingly  increased. 

The  beds  of  surface  streams  must  be  relatively  impervious,  for  if  they 
were  freely  percolated  by  water,  there  would  soon  be  no  visible  flow. 
Where  a  stream's  bed  is  is  partially  porous,  much  of  the  water  sinks  to  the 
first  impervious  stratum  and  there  forms  an  invisible  stream  called  the 
underflow  which  often  contains  more  water  than  its  parent  overhead. 
Where  a  stream  has  not  naturally  a  channel  of  impervious  rock  or  clay, 
the  tendency  is  for  it  to  stop  the  pores  of  a  sandy  or  other  pervious  bed 
with  sediment;  especially  is  this  so  in  alluvial  valleys  like  that  of  Fig.  19, 
where  there  might  be  no  visible  stream  at  all  had  the  river  at  r  not  a 
clay-coated  bottom. 

For  our  drainage  problem  of  Fig.  19,  we  have  now  discussed  how  to 


50  MINING    WITHOUT    TIMBER 

ascertain  the  area  of  watershed  A,  the  depth  of  rainfall  D,  the  evapora- 
tive and  run-off  factors  e  and  r,  and  by  substituting  these  values  in  equa- 
tion (4)  we  have  the  percolation  P.  The  result  from  solving  Equation 
(4)  can  be  compared  with  the  following  table  which  gives  for  the  percen- 
tage of  total  rainfall  percolating  various  surfaces: 

sand  =  60  to  70 
chalk  or  gravelly  loam  =  35 
sandstone  =  25 
limestone  =  15 
clay  or  granite  =  15  and  less 

We  do  not  have  to  provide  at  s  for  the  drainage  of  volume  P,  but  only 
for  that  portion  of  it  which  is  not  drained  off  elsewhere,  does  not  reascend 
to  evaporate  at  the  surface  or  is  not  held  in  the  pores  of  the  subsoil. 
The  drainage  elsewhere  would  be  nil  in  a  lake  basin  with  impervious 
bottom,  but  in  the  usual  self-draining  basin  it  would  constantly  tend  to 
lower  the  water  table. 

The  lake-basin  condition  is  well  exemplified  at  both  Bisbee  and 
Tombstone,  Arizona.  These  camps  lying  in  the  Mule  Mountains,  where 
the  annual  rainfall  is  under  12  in.  and  the  evaporative  factor  large,  would 
be  casually  reckoned  as  having  dry  mines,  but  the  very  opposite  is  the 
case.  The  ore  bodies  in  each  camp  are  found  in  limestone  and  shale 
beds  which  are  so  folded  as  to  form,  with  the  adjoining  intrusive  rocks, 
an  impervious  basin  which  catches  all  the  rain  percolating  the  surface  over 
a  large  watershed.  The  present  water  in  the  basins  represents  the 
accumulation  of  years,  so  to  lower  it  has  taken  more  pumping  than 
would  be  necessary  in  a  very  wet  valley  whose  ground  water  was  depend- 
ent solely  on  current  rainfall.  At  Bisbee  in  1906  the  water  level  had 
been  permanently  lowered,  for  three  years'  pumping  by  several  companies 
had  reduced  the  inflow  at  the  Calumet  and  Pittsburg  shaft  from 
3000  to  1500  gal.  per  min.;  but  at  Tombstone  in  1911,  where  nine  years 
of  pumping  of  the  Contention  shaft  had  little  affected  the  original  flow 
of  3000  gal.  per  min.,  it  was  deemed  unprofitable  to  struggle  further  and 
the  pumps  were  pulled.  The  excess  of  water  in  the  Tombstone  basin 
probably  comes  from  an  adjoining  watershed  through  underground 
channels. 

The  loss  of  ground  water  by  evaporation  increases  with  a  damp  soil 
and  a  high  water  table.  Consequently,  in  self-draining  ground  the 
evaporation  is  greater  if  the  rainfall  is  evenly  rather  than  sporadically 
distributed.  Evaporation,  however,  usually  affects  the  water  table  only 
slightly  as  compared  with  the  capacity  of  the  formation  for  the  storage, 
surrender  and  passage  of  water. 

In  Fig.  19,  the  section  of  the  water-storage  area  between  wt  and  bed- 
rock is  not  the  whole  area  wmnt,  but  this  area  multiplied  by  the  factor 


PRINCIPLES    OF    MINE    DRAINAGE  51 

for  "voids"  or  the  proportion  of  intergranular  spaces  in  the  formation. 
The  void  factor  depends  less  on  the  size  of  rock  grains  than  on  their  uni- 
formity, and  varies  from  0.2  to  0.5.  Yet  another  item  must  be  included, 
in  estimating  the  quantity  of  water  that  must  be  drained  to  lower  wt,  and 
that  is  the  factor  for  surrender  or  yield  which  depends  on  the  capillarity 
or  fineness  of  the  grain  of  the  formation.  The  yield  factor  is  almost  nil 
for  clay,  0.5  to  0.6  for  porous  soil,  0.6  to  0.7  for  sand,  and  nearly  1.0  for 
clean  gravel  or  boulders.  A  low  yield  factor  means  not  only  the  reten- 
tion of  rainfall  in  a  porous  formation  until  it  is  saturated,  but  a  long  delay 
before  a  heavy  shower  begins  to  be  noticed  underground.  From  the 
above,  if 

W=  volume  in   cu.   ft.   of  water-bearing  formation  tributary  to 

sump  s 
W'=  volume  in  cu.  ft.  of  water  the  water-bearing  formation  yields 

tributary  to  sump  s 
x  =  factor  for  voids  in  formation 
y  =  factor  for  water-yield  of  formation 

then  W'=xyW.  (5) 

To  free  our  placer  ab  from  water,  it  will  not  be  necessary  to  lower  the 
water  table  to  the  profile  wmnt,  but  only  to  the  profile  whabgt  where  who, 
and  bgt  are  the  profiles  of  the  hydraulic  gradient  toward  the  sump  s.  The 
hydraulic  gradient  increases  with  the  fineness  of  grain,  though  very  small 
in  gravel,  it  is  30  to  50  ft.  per  mile  in  sand  and  in  a  large  basin,  it  would 
thus  considerably  decrease  the  volume  of  water  tributary  to  sump  s. 

The  hydraulic  gradient  for  a  given  formation  can  be  directly  measured 
by  digging  two  wells  in  the  same  line  of  water  drainage,  at  some  distance 
apart,  and  then  recording  their  water  levels.     The  hydraulic  gradient 
will  then  be  the  difference  of  water  level  divided  by  the  distance  between  ' 
the  wells. 

Hazen1  gives  as  a  formula  for  the  velocity  of  passage  of  ground  water 
V  =  KD*S  (6) 

where  V  =  velocity  in  ft.  per  sec.  of  flow  through  ground  pores 
.£  =  0.29  (a  constant) 

D  =  diameter  in  mm.  of  sand  grain  "effective"  (i.e.,  90  per  cent,  of 
grains  must  be  larger  than  D).  Formula  (6)  is  inapplicable 
when  d  is  less  than  3 

/S  =  sine  of  slope  of  the  hydraulic  gradient 

Then  if  B  =  area  in  sq.  ft.  of  a  vertical  surface  enclosing  mine 
openings  extending  from  water  surface  in  pump  to  water 
table.  Height  of  surface  B  should  be  small  for  use  of 
formula  (6) 

/=  volume  of  water  in  cu.  ft.  entering  placer  per  sq.  ft.  of  area  B 
F  =  total  volume  in  cu.  ft.  entering  placer  over  total  area  B 
kf  =  fractional  factor  for  voids  in  walls  of  area  B 
i  Massachusetts  State,  Reports  on  Water  Supply. 


52  MINING    WITHOUT    TIMBER 

It  is  evident  that 


Substitute  for  V  from  Equation  (6)  and 

F=kK'BD*S  (7) 

In  practice  the  possibility  of  keeping  the  placer  dry  enough  to  permit 
miners  to  work  would  of  course  depend  not  only  on  the  means  of  drainage 
available  to  keep  sump  s  clear,  but  also  on  /  or  the  rate  of  inflow  at  the 
mining  face.  As  /  increased  beyond  a  certain  figure,  the  miners  would 
find  themselves  working  in  a  heavy  spray  and  standing  in  a  gurgling  pond. 
In  such  a  case,  unless  the  inflow  could  be  controlled  by  a  cofferdam,  sub- 
aqueous mining  would  have  to  be  resorted  to. 

Case  II.  Deposits  in  Stratified  Rock.  —  An  example  of  this  case  is 
shown  in  Fig.  20,  a  cross-section  of  a  coal  seam  A  in  a  synclinal  basin. 
Beneath  the  coal  is  a  thin  layer  of  clay  B  resting  on  a  sandstone  C,  and 


FIG.  20. — Drainage  in  stratified  rock. 

above  it  are  strata  of  sandstone  D,  shale  G,  and  limestone  H.  Along  the 
surface  runs  a  river  r  over  a  valley-filling  of  alluvial  soil.  Then  the  perco- 
lation into  a  coal  mine  at  A  will  depend  not  only  on  the  coal  itself  whose 
bedding  and  joint  planes  may  be  somewhat  permeable,  but  on  the  nature 
of  the  adjoining  rocks. 

Clay  and  shale  are  not  only  relatively  impermeable  but  plastic,  and 
tend  to  close  tm  any  openings  made  inyaving  aorogenic  movements. 
Sandstones  vary  in  their  structure,  some  h  duebh  texture  as  porous  as 
free  sand,  while  the  grains  of  others  are  closely  cemented  and  almost 
impermeable.  Limestones,  especially  if  dolomitic,  abound  in  irregular 
channels  and  pot-holes,  often  large  enough  to  contain  underground  rivers 
or  ponds.  Should  the  rocks  of  Fig.  20  be  subjected  to  metamorphism, 
their  permeability  would  be  much  diminished,  or  perhaps  entirely  de- 
stroyed, as  pores  and  bedding  planes  were  obscured,  until  we  approached 
as  the  limit  the  massive  formation  of  Case  III.  Clay  and  shale,  when 
metamorphosed,  become  dense  and  strong  slate  or  schist,  sandstone  solidi- 
fies into  impermeable  quartzite,  and  limestone  changes  into  crystalline 
marble. 

From  these  considerations,  it  can  be  seen  that  the  stratum  of  shale 
at  G,  provided  it  has  not  been  pierced  by  erogenic  movements  or  human 
hand,  acts  as  a  screen  to  keep  out  any  water  which  may  percolate  into 
the  limestone  from  the  watershed  ef.  As  the  strata  outcrop,  however, 


PRINCIPLES    OF    MINE    DRAINAGE  53 

beyond  the  summits  e  and  /  of  the  synclinal  basin,  the  coal  seam  will  be 
exposed  to  percolation  from  watersheds  ec  and/c'. 

As  long  as  the  impermeable  clay  floor  B  of  the  coal  is  uncracked,  the 
watersheds  contributory  to  the  coal  seam  will  extend  only  from  e  to  b  and 
from  /  to  61  and  not  all  of  their  percolation  will  reach  the  coal,  because 
the  shale  stratum  G  will  seal  off  any  surface  water  that  may  enter  the 
limestone  layer  H  between  the  crests  e  and  /  and  the  roof  of  G.  Should 
the  floor  B  be  cracked  or  feathered  out  in  places,  it  may  be  serious  from 
a  drainage  standpoint,  for  the  coal  seam  will  then  be  open  to  a  flood  from 
the  hydrostatic  water  in  sandstone  (7.  Thus  in  the  rock  formation  of 
Fig.  20,  the  ground  water  would  not  occur  in  a  connected  body  as  in  Fig. 
19,  but  each  porous  zone  would  contain  its  own  pool  separated  from  the 
others  by  an  impermeable  stratum.  The  equivalent  of  the  water  table 
wt  of  Fig.  19  would  be  found  here  in  the  limestone  H,  but  it  would  cir- 
culate there  in  irregular  open  channels  instead  of  in  intergranular  pores. 
The  contributory  watersheds  having  been  thus  measured,  we  have  only 
then  to  gather  the  other  meteorological  and  physical  constants,  as  ex- 
plained for  Case  I,  in  order  to  solve  Equations  (1)  to  (7)  for  the  drainage 
of  Fig.  20. 

Case  HI.  Deposits  in  Massive  Rock. — In  Fig.  21,  let  cd  be  the  cross- 
section  through  a  fissure  vein  in  massive  rock,  which  is  either  of  igneous 


FIG.  21. — Drainage  in  massive  rock. 

origin  or  so  metamorphosed  that  its  sedimentary  pores  and  bedding  planes 
are  practically  obliterated.  This  formation,  then,  instead  of  being  quite 
porous  like  that  of  Case  I  or  irregularly  porous  like  that  of  Case  II,  is  in 
its  original  condition,  more  or  less  impermeable,  but  in  mining  regions 
it  has  usually  been  so  cracked  by  earth  movements  as  to  abound  in  open- 
ings which  grade  from  wide  fissures,  both  long  and  deep,  to  such  minute 
fracture  planes  as  those  of  the  Bingham  copper  porphyry  which  scarcely 
pass  seepage  water.  When  rainfall  can  only  percolate  the  surface  of 
Fig.  21,  through  irregularly  spaced  crevices  or  joints  instead  of  through 
a  porous  zone,  there  can  be  nothing  like  a  general  ground  water  level 
except  within  areas  whose  crevices  are  all  connected.  Thus  each  crevice 
system  has  a  height  of  water  table  varying  according  to  the  size  and  nature 


54  MIXING    WITHOUT    TIMBER 

of  its  contributory  watershed.  The  mineral  veins  themselves  have  often 
trunk  channels  along  their  walls  which  receive  water  from  numerous 
branch  cracks  and  fissures. 

The  watershed  tributary  to  vein  cd  of  Fig.  21  will  not  extend  laterally 
from  e  to  /  as  in  Case  I  unless  all  the  intermediate  fissure  systems  lead  to 
the  vein,  but  it  may  cover  a  much  wider  area  owing  to  the  possible 
juncture  of  subterranean  streams  with  cd,  which  streams  in  mountainous 
regions  may  be  under  a  high  hydrostatic  head.  In  fact,  only  the  map- 
ping of  the  region's  underground  water  channels,  and  this  could  seldom  be 
done  except  in  an  extensively  developed  district,  would  enable  an  engi- 
neer to  satisfactorily  solve  Equations  (1)  to  (7)  as  in  the  two  previous 
cases.  In  mining  cd,  care  would  have  to  be  taken  on  the  hanging  side, 
for  by  the  tapping  of  natural  blind  crevices  or  by  allowing  the  hangwall 
to  move  and  crack,  the  river  r  might  be  precipitated  into  the  workings. 

It  is  probable  that  some  of  the  hot  water  found  in  mining  such  igneous 
formations  as  the  Comstock  lode  comes,  not  from  rainfall,  but  directly 
from  the  occluded  moisture  of  cooling  magmas.  According  to  the  nebu- 
lar hypothesis,  all  surface  water  had  originally  an  igneous  origin.  The 
miner  who  operates  in  a  region  of  magmatic  water  cannot  estimate  its 
quantity  beforehand,  as  in  the  case  of  meteoric  inflows,  but  must  simply 
handle  it  as  it  appears. 

As  mines  get  deeper  and  rock  pressures  become  greater,  fissures  and 
other  open  spaces  tend  to  close  up  and  long  before  the  bottom  of  the  zone 
of  crust  fracture  is  reached,  at  a  depth  of  somewhere  around  4  miles,  there 
is  little  free  water  in  the  rocks.  At  the  Calumet  and  Hecla  Copper  mine, 
Mich.,  in  a  conglomerate  lode  bedded  between  amygdaloids,  the  maxi- 
mum water  flow  is  at  1800  ft.  along  the  38  deg.  dip,  while  at  3000  ft.  the 
water  flow  is  insufficient  even  to  supply  the  drills. 

CONTROL  OF  WATER 

This  topic  naturally  divides  itself  into  surface  and  underground 
control. 

Surface  Diversion. — It  is  much  better  to  keep  water  out  of  a  mine  than  to 
use  the  most  approved  method  of  drainage  after  its  unnecessary  entrance. 
Surface  run-off  is  kept  out  of  a  mine  ditching  around  shafts  and  vein  out- 
crops as  C  in  Fig.  21.  Often  it  is  best  to  refrain  from  stoping  a  vein  out 
quite  up  to  the  surface  in  order  to  keep  rain  out  of  the  workings.  A 
stream  above  the  mine,  which  seeps  badly  into  its  bed  or  whose  bottom 
may  be  cracked  by  caving  operations,  can  often  be  diverted  to  another 
channel  or  carried  in  a  flume  over  the  dangerous  stretch. 

Underground  Diversion.— In  Fig.  20  the  penetration  of  the  imperme- 
able shale  layer  G  by  the  shaft  A  A'  will  eventually  drain  the  wet  lime- 
stone layer  H  into  the  mine  at  A  unless  some  precaution  is  taken.  Two 


PRINCIPLES    OF    MINE    DRAINAGE 


55 


remedies  suggest  themselves,  the  first,  a  concrete  shaft-lining  from  the 
surface  down  to  a  sealed  footing  in  the  shale;  the  second,  the  usual  pervi- 
ous shaft-lining  provided  with  a  water  ditch  or  "ring"  around  the  shaft, 
in  the  roof  of  the  shale,  which  catches  the  water  from  above  and  leads  it 
to  a  sump,  which  has  means  for  drainage,  on  the  same  level. 

In  ore  deposits  in  hilly  regions,  an  impervious  floor  sometimes  has 
below  it  a  sandstone  or  other  porous  stratum  which  dips  toward  an  out- 
crop, on  a  hillside  at  a  lower  level,  and  is  thus  self-draining.  In  such  a 
case,  diamond-drill  holes  or  a  winze  through  the  floor  of  the  mine  sump 
into  the  porous  stratum  will  effectively  drain  the  workings. 

Natural  Dams. — Rock  barriers  are  highly  useful  in  the  control  of 
water  in  mines.  As  already  explained,  where  tight  strata  cut  off  the 
mine  from  wet  formations,  such  natural  seals  should  be  left  undisturbed 
if  possible.  Pillars  of  mineral  are  often  left  between  adjoining  mines  to 
keep  their  water  systems  distinct,  and  in  many  states  a  barrier  about 
50  ft.  wide  must  be  left  unminad  around  the  boundary  of  coal  properties. 

The  Lehigh  Valley  Coal  Co.  is  now  mining,  near  Hazleton,  Pa.,  a 
synclinal  trough  containing  parallel  anthracite  seams  which  extends  for 
several  miles  and  dips  for  about  3000  ft.  vertically  in  that  distance.  The 
trough  has  been  divided  into  three  drainage  basins  by  leaving  a  trans- 
versal barrier  pillar  of  coal,  100  ft.  wide,  below  each.  The  barriers  are  at 
altitudes  of  1084,  1250,  and  4000  ft.  respectively  and  each  has  its  own 
unwatering  system.  Each  barrier  is  pierced  by 
boreholes  lined  with  pipes  whose  valves  can  be 
opened  to  drain  the  basin  above  into  the  one  below 
in  case  of  an  emergency. 

It  is  often  necessary  to  penetrate  water  barriers 
in  order  to  drain  old  mines,  and  where  the  dammed- 
up  water  is  under  a  high  head  it  is  best  tapped  by 
drilling.  Boring  long  holes  for  tapping  can  be 
done  in  any  direction  by  a  diamond  drill.  A 
customary  safe-guard  against  heavy  pressure  is  to 
bore  the  first  few  feet  of  the  hole  large  enough  for 

a  pipe  lining  ck,  Fig.  22,  whose  exterior  is  made  to  fit  the  rock  tightly 
by  a  packing  of  lamp-wick,  wound  spirally,  or  of  cement.  When  it  is 
necessary  to  regulate  the  flow  from  the  hole,  a  valve  v  is  put  on  the 
lining  pipe  whose  end  must  then  be  anchored  to  posts  p.  Such  a  valve 
on  a  drill-hole  flow,  small  anyhow,  allows  the  cautious  emptying  of  old 
workings  where  a  sudden  release  of  water  might  damage  the  shaft  or 
other  important  pillars. 

For  short  distances,  tapping  can  be  well  performed  with  a  percussive 
drill  and  a  typical  recent  case  can  be  cited  at  the  Iron  Mt.  Mine,  Montana, 
where  the  new  drainage  adit  was  connected  with  the  old  shaft-workings 
which  contained  some  100,000,000  gal.  of  water  under  a  head  of  900  ft. 


FIG.    22. — Tapping  sump  a 
Iron  Mt.  Mine,  Montana. 


56 


MINING    WITHOUT   TIMBER 


When  the  face  of  the  tunnel  to,  Fig.  22,  arrived  near  the  old  shaft  s,  a 
6x7-ft.  cross-cut  tc  was  run  for  30  ft.,  parallel  to  the  shaft  station,  and 
from  c  a  taildrift  was  carried  back  for  30  ft.  to  d.  Next  a  3-in.  percussive 
drill  was  set  up  at  c  and  a  hole  drilled  in  for  10  ft.  to  admit  a  4-in.  pipe 
lining  ck  which  was  then  well  cemented  and  anchored  in.  Drilling  was 
proceeded  beyond  the  lining  with  a  1  3/4-in.  bit  when,  at  e,  23  ft.  from 
the  collar,  the  point  holed  through.  On  loosening  the  chuck,  the  bit 
was  shot  back  by  the  water  pressure  against  the  end  c,  and  was  followed 
by  a  swift  stream  of  water,  but  as  a  low  dam  had  been  erected  across  the 
crosscut  at  t,  the  men  climbed  over  it  and  safely  reached  the  adit  mouth, 
a  mile  and  a  half  distant. 

Artificial  Dams. — Many  dams  are  built  underground:  for  making 
sumps  out  of  old  headings  or  stopes;  for  regulating  the  flow  to  the  pumps; 
for  isolating  the  water  of  abandoned  workings;  and  for  confining  water  to 


Section  x-y 
Fid.  23. — Artificial  dama. 

certain  localities,  as  in  the  case  of  flooding  mine  fires  or  of  filling  seams 
by  the  flushing  system.  Mine  dams  differ  from  those  on  the  surface  in 
the  fact  that  they  often  stop  openings  of  small  height  relative  to  the 
pressure  of  water  to  be  sustained.  In  such  cases,  mine  dams  must  have  a 
solid  footing  all  around  their  periphery  instead  of  just  at  the  base  and 
sides  like  a  river  dam.  Favorite  materials  for  dams  are  wood,  brick, 
stone  and  concrete. 

A  diverting  dam  whose  crest  is  higher  than  the  water  surface  it  sustains 
can  be  built  light  and  like  a  surface  structure;  but  precautions  must  be 
taken  to  successfully  sustain  a  high  water-head  (which  causes  a  pressure  of 
0.434  Ib.  per  sq.  in.  for  each  foot  of  height),  and  the  arch  is  a  favorite  form 
for  this  purpose.  Fig.  23  shows  a  composite  plan  and  section  of  a 
dam,  to  back  up  water  at  y  across  a  heading  or  shaft,  which  is  made  of 
two  arches,  ab  and  cd,  with  a  filling  between  of  puddled  clay  or  concrete. 
It  will  be  noticed  that  the  heading  walls  are  cut  out  to  give  indented  skew- 
backs  for  the  arches  except  at  e'f  and  g'  hf  where  a  plastic  roof  and  floor 


PRINCIPLES    OF    MINE    DRAINAGE  57 

might  make  indentation  unnecessary  if  the  swelling  wood  construction, 
to  be  described  later,  were  used. 

Both  drain-pipe  at  m  and  air-escape  at  n  are  provided  with  valves  and 
sealed  tightly  in  the  structure.  The  manhole  pipe  xy  is  anchored  to  the 
dam  and  is  as  essential  during  construction  as  afterward.  With 
moderate  pressure  one  arc}},  like  eh,  is  enough;  and  to  build  it  of  wood, 
each  piece  should  be  the  length  of  eh  and  tapered  wedge-shape  like  an 
arch  stone.  A  tight  joint  can  be  made  between  wood  and  walls  by 
tarred  felt  and  small  wedges,  and  the  pipes  can  be  sealed  in  with  wedges. 
When  of  masonry,  the  arches  are  laid  over  wooden  centers,  the  under  one 
of  which  is  left  in  permanently  if  the  dam  is  across  a  shaft.  Masonry 
dams  are  kept  tight  by  a  concrete  or  clay  backing,  and  as  the  latter 
needs  to  be  confined  under  heavy  pressure,  the  double  arches  of  Fig.  23, 
with  clay  between,  are  then  especially  suitable.  Flat  wooden  dams 
are  often  used  and  usually  they  are  held  by  posts  with  ends  hitched  into 
the  walls.  The  wooden  lining  is  made  of  several  layers  of  planks  and, 
with  walls  too  soft  for  its  support  by  posts  at  intervals  in  hitches,  the 
lining  itself  may  be  extended  into  brick-lined  hitches  cut  in  the  walls, 
and  its  central  portion  be  backed  by  a  timber  set  whose  battered  posts  are 
set  in  the  direction  of  pressure  and  rest  in  hitches  cut  in  the  heading's 
walls. 

At  the  Chapin  iron  mine,  Michigan,  dams  have  been  helpful  in  the  con- 
trol of  a  big  inflow  of  water.  The  Chapin  ore-body  is  a  hematite  lense 
appearing  in  cross-section  about  like  cd  in  Fig.  21.  It  is  enclosed  by  slate 
walls  but  has  an  extensive  dolomite  formation  about  100  yards'  distant  on 
the  hanging  side.  The  author  found  on  his  visit  in  1908  that  the  flow  at 
the  1000-ft.  level  had  not  been  appreciably  lessened  in  spite  of  pumping 
2000  to  3000  gal.  per  min.  for  the  previous  seven  years.  The  water 
proceeds  from  channels,  in  the  dolomite  hangwall,  which  are  believed  to 
connect  with  two  small  lakes  several  miles  to  the  northeast.  If  the 
originally  impermeable  slate  hangwall,  that  cut  off  the  ore  from  the 
water-bearing  dolomite,  had  not  been  cracked  for  over  300  ft.  from  the 
surface  (by  the  caving  operations),  the  drainage  problem  could  easily 
have  been  solved  by  keeping  shafts  and  cross-cuts  entirely  in  foot-wall. 

The  No.  2  Hamilton,  vertical  shaft,  then  used  for  pumping,  had  been 
sunk  in  the  hanging  dolomite  and  great  difficulty  had  been  encountered 
in  driving  the  1000-ft.  and  lower  cross-cuts  because  of  the  water  crevices 
encountered.  In  starting  the  1000-ft.  cross-cut  a  compound  station  pump 
was  first  installed;  but,  nevertheless,the  first  water  crevice  struck  had  to 
be  dammed  with  masonry,  and  the  pressure  gauge  showed  a  static  head 
there  from  a  water  level  within  300  ft.  of  the  surface.  Next,  a  branch 
drift,  some  distance  back  from  dam  No.  1,  was  begun;  but  this  also 
struck  a  crevice  and  had  to  be  dammed. 

A  second  branch  drift  was  then  started  and  dammed  (after  only  a 


58  MINING    WITHOUT    TIMBER 

short  advance)  and  a  second  compound  pump  installed  at  the  station. 
This  last  dam  was  fitted  with  a  water-tight  iron  door  opening  outward, 
so  when  drifting  was  continued  beyond  it  (to  make  a  chamber  for  diamond 
drilling)  the  excavated  earth  could  be  passed  back  in  boxes.  With  the 
diamond  drill,  the  space  yet  to  transverse  to  reach  the  vein  was  searched 
for  a  cross-cut  opening  free  from  crevices;  b^ut  as  none  was  found  the 
cross-cut  had  finally  to  be  finished  anyhow  by  the  aid  of  strenuous 
pumping. 

DAMMING  BY  DEPOSITION 

E.  B.  Kirby  has  devised  a  method  (U.  S.  Pat.  No.  900,683)  of  sealing 
the  rock  crevices  of  mine  workings  by  the  deposition  therein  of  sedi- 
ment. Finely  divided  clay  is  preferable  but  other  materials  may  be 
used  such  as  sand,  mill  tailing  or  slime,  cement,  saw-dust,  horse  manure, 
chopped  hay  or  fiber.  The  injection  of  the  water  bearing  this  material 
in  suspension  may  be  made  by  force-pumps,  or  by  stand-pipes  extending 
far  enough  toward  the  surface  to  furnish  the  necessary  pressure. 

The  suspended  particles,  when  put  in  a  cavity  containing  water  in 
motion  toward  exits  in  the  mine,  are  seized  and  carried  toward  such 
exits,  settling,  accumulating  in,  and  choking  at  various  points  the  con- 
tributory passages.  The  moving  currents  automatically  select  those 
passages  which  are  discharging  water  into  the  mine  and  require  sealing; 
they  disregard  other  passages  because  the  water  is  not  in  motion  in 
them.  The  choking  which  occurs  in  the  outflowing  passages  is  gradual 
and  at  those  most  favorable  points  where  the  passages  are  smallest  and 
the  flow  most  diffused.  In  fact  large  passages  cannot  be  thus  choked 
but  must  by  dammed. 

When  the  flowing  passages  are  choked  the  process  ceases  even  though 
other  passages  are  still  open.  If  by  the  choking  of  one  or  more  passages 
the  current  is  deflected  to  others,  the  deposition  is  there  automatically 
resumed.  At  any  choked  locality  the  water  pressure  holds  the  choking 
particles  firmly  in  place  and  produces  a  perfect  seal  by  shutting  off  the 
threads  of  water  in  every  contributing  passage. 

Adits. — These  are  tunnels  run  in  from  a  low  surface  point  to  drain 
underground  workings.  In  Fig.  21,  it  is  evident  that  the  adit  ad  would 
drain  all  of  vein  cd  above  level  d  and  that  adit  bk  would  drain  everything 
above  level  k.  The  water-bearing  fissure  mn  cuts  the  vein  at  n,  and  to 
drive  an  adit  at  the  level  n  would  obviously  be  impractical  with  the  given 
topography;  but  nothing  would  hinder  the  extension  of  adit  bk  to  the 
fissure,  and  the  running  of  a  drift  g  along  its  footwall,  as  far  as  necessary, 
to  intercept  all  the  water  drainage  into  the  mine  at  n.  This  scheme  was 
employed  to  supplement  the  lower  adit  of  the  Horn  Silver  mine  in  Utah. 

The  following  remarks  apply  to  all  drainways,  whether  adits  proper, 
debouching  at  the  surface,  or  merely  interior  tunnels  emptying  into  a 


PRINCIPLES    OF    MINE    DRAINAGE 


59 


sump.     The  minimum  grade  for  long  modern  adits  with  unlined  ditches 
is  1/4  of  1  per  cent.     The  carrying  power  of  water  channels  can  be  thus 
estimated: 
If 

Sine  of  slope  of  hydraulic  gradient  of  water  flowing  in 

channel  =S 

Area  of  cross-section  of  water  flowing  in  channel  =a  sq.  ft. 

Velocity  per  sec.  of  water  flowing  in  channel  =v  ft. 

Wetted  perimeter  of  the  containing  surface  of  channel  =  P  ft. 

Constant,  increasing  with  smoothness  of  containing  sur- 

face of  channel  —c 

quantity  of  water  flowing  per  sec.  in  channel  =q  cu.  ft. 
then  from  Merriman's  "  Hydraulics  " 


but  q  = 


,'aS 
hence  5  =  ac-l  — 


(9) 


In  those  cases  where  adits  are  only  to  be  used  for  drainage,  a  circular 
section  is  often  preferable;  because  it  carries,  when  running  half  full  or 
more,  the  most  water  for  a  given  volume  of  exca- 
vation; is  stable  against  external  pressure;  and  is 
readily  adaptable  to  masonry  lining,  which,  being 
smoother  than  wooden  sets,  gives  a  larger  value  for 
c  in  formula  (9),  and  consequently  passes  more 
water.  Where  the  water  deposits  sediment,  the 
egg-shaped  section  of  Fig.  17  (4),  used  for  sewers, 
best  enables  a  uniform  carrying  power  to  be  main- 
tained as  the  height  of-  water  fluctuates. 

When  adits  serve  for  haulage  as  well  as  drain- 
age, the  economical  shape  is  usually  oval  or  rec- 
tangular. The  oval  shape  is  best  for  weaker  walls 
with  external  pressure  mainly  vertical,  and  it  can 
easily  be  lined,  where  necessary,  by  masonry.  The 
rectangular  shape  is  common  where  the  adit  fol- 
lows some  flat  stratum  like  a  coal  seam  with  a  strong  roof,  that  will  stand 
without  arching;  or  where  most  of  the  length  has  to  be  supported  by 
timber  or  metal  sets  which  are  ill  adapted  to  curved  sections. 

A  compromise  section  for  an  adit  is  shown  in  Fig.  24  (a)  with  a  self- 
sustaining  arched  roof  and  a  flat  bottom  to  give  a  cheap  footing  for  the 
track  ties.  With  a  moderate  amount  of  water  it  can  be  carried  in  a  side 
ditch  which  is  easier  to  watch  and  clean  than  one  under  the  track.  Where 
the  adit  of  Fig.  24  (a)  is  in  a  narrow  vein  of  width  from/'  tod',  the  ditch 


h          h 
FIG.    24. — Adit  sections. 


60  MINING    WITHOUT    TIMBER 

is  best  placed  along  that  side  whose  cutting-out,  to  give  space  for  the 
adit,  will  admit  the  least  water  from  the  walls.  The  tightness  of  the 
ditch's  bottom  rock  against  seepage  should  also  be  considered,  if 
there  are  to  be  workings  underneath  it,  and  sometimes  a  wooden  or 
concrete  lining  may  be  necessary  as  commonly  it  is  for  sumps.  Where 
the  adit  can  be  placed  between  vein  walls  as  ef  and  ed  without  cutting 
them,  it  is  usually  best  to  have  the  main  ditch  along  the  footwall  at  c; 
and  connect  it,  if  necessary,  by  cross  ditches  to  an  auxiliary  ditch  along 
the  hangwall  at  e'. 

Both  ditches  and  sumps  should  be  covered  in  hot  mines  like  those 
of  the  Comstock  lode  in  order  to  prevent  any  unnecessary  humidifying 
of  the  air. 

The  new  Roosevelt  adit  at  Cripple  Creek,  Colo.,  will  be  over  3  miles 
long  and  used  only  for  drainage.  This  gold  mining  district  lies  in  an 
igneous  formation,  and  as  it  occupies  an  area  of  about  8  sq.  miles,  it  is 
estimated  that  each  foot  in  height  of  its  ground  water  means  60,000,000 
gal.  of  water.  The  adit  was  started  with  a  section  like  Fig.  24  (a),  10  ft. 
high  and  6  ft.  wide,  but  it  has  been  changed  to  one  6  ft.  high  by  10  ft.  wide 
to  give  space  for  a  curved  ditch,  6  ft.  wide  by  3  ft.  deep,  and  a  narrow 
track  along  one  wall. 

Where  side  ditches  are  inconvenient  or  inadequate,  they  can  be  re- 
placed or  supplemented  by  a  central  ditch  nc'  (shown  dotted  in  Fig.  24  (a)) 
cut  under  the  rails.  For  heavy  flows,  however,  the  whole  bottom  of  the 
adit  may  be  utilized.  In  that  case,  it  should  be  cut  round,  as  ghh'k  in 
Fig.  24  (b),  in  order  to  obtain  the  cheapest  rock  breaking  and  the  maxi- 
mum carrying  power  for  a  given  sectional  area;  unless  a  flat-bedded  forma- 
tion makes  the  excavation  of  the  larger  square  areagg'k'k  j  ust  as  economical. 
The  track  ties  may  be  spiked  to  stringers  which  are  set  on  posts  or  brick 
piers,  h  and  h',  of  sufficient  height  to  keep  the  ties  above  the  high-water 
flow.  In  double  track  adits,  three  rows  of  stringers  on  piers  are  sufficient 
if  long  ties  are  used.  Where  the  track  is  far  above  the  rock  bottom  and 
the  adit  is  narrow,  cross  beams  like  gk,  hitched  into  the  walls,  may  be  the 
cheapest  supports  for  the  stringers. 

Adits  are  especially  advantageous  in  mountainous  regions  of  steep 
slopes,  where  a  great  height  can  be  drained  with  a  short  adit.  The  only 
drainage  expense  with  adits  is  for  interest  and  maintenance,  and  if  well- 
constructed,  they  are  not  subject  to  the  breakdowns  of  mechanical  appa- 
ratus at  critical  moments.  When  the  adit  mouth  is  some  distance  higher 
than  the  stream  into  which  it  drains,  the  escaping  water  can  be  effectively 
utilized  for  power.  In  a  wet  district  of  large  producing  mines  whose 
drainways  can  easily  be  connected,  it  is  often  advisable  to  drive  a  very 
long  adit  for  general  drainage. 

Notable  among  such  modern  American  adits  are,  in  Colorado,  the 
Roosevelt  and  the  5-mile  Newhouse  at  Idaho  Springs;  and  in  the  an- 


PRINCIPLES    OF    MINE    DRAINAGE  61 

thracite  region  of  Pennsylvania,  the  5-mile  Jeddo- Basin  in  Luzerne  Co., 
the  1-mile  Oneida  in  Schuylkill  Co.,  and  the  1  1/2-mile  Lausanne  near 
Mauch  Chunk.  The  last  named  drains  13  miles  of  underground  tunnels 
and  14  different  collieries. 

Siphons. — In  mining  flat  coal  and  other  seams,  convex  rolls  often 
occur  in  the  floor  of  the  gangways  which  dam  up  the  drainage.  It  is 
feasible  to  pass  a  low  roll  by  deepening  the  water  ditch;  but  a  high  roll, 
unless  it  is  advantageous  to  also  cut  the  whole  gangway  through  it  to 
obtain  a  uniform  haulage  grade,  is  often  better  surmounted  by  a  siphon. 
A  siphon  consists  of  a  vertically-curved  pipe  with  both  ends  set  in  sumps, 
of  which  the  outlet  sump  must  have  the  lowest  water-level. 

Mine  siphons  are  usually  made  from  welded  iron  pipe  and  water  can 
be  carried  horizontally  in  them  for  considerable  distances  provided  they 
are  tight!  The  limit  of  vertical  lift,  from  surface  of  intake  to  highest 
point  on  the  pipe  of  any  siphon,  is  the  height  of  the  water  barometer 
minus  the  total  loss  of  water  head,  due  to  internal  friction,  etc.,  in  the 
siphon  itself.  This  limit  is  usually  below  26  ft.  Several  rolls  can  be 
passed  by  one  siphon  if  escape  valves  for  air  are  put  on  the  pipe  at  the  high 
point  of  each  vertical  bend.  It  is  also  possible  to  drain  several  sumps  or 
"  swamps  "  along  a  gangway  with  one  siphon,  by  running  a  branch  pipe, 
with  a  valve  on  its  end,  down  from  the  main  siphon  into  each  swamp. 
A  siphon  is  best  rigged  with  a  valve  at  inlet  and  outlet;  and  with  its 
highest  point  joined  by  a  small  pipe,  with  valve,  to  a  water-barrel  from 
which  it  can  easily  be  filled  before  a  run. 

MECHANICAL  'UNWATERING 

Apparatus  for  mechanical  drainage  can  be  grouped  into  two  classes. 
First,  those  moving  water  in  buckets,  and  second,  those  moving  water 
through  pipes.  In  the  first  class,  water-cars  are  moved  horizontally  by 
the  same  tractors  as  ore-cars,  while  tanks  or  kibbles  are  hoisted  in  shafts 
or  slopes  by  similar  engines  to  those  used  for  hoisting  ore  in  skips.  The 
second  class  includes  all  types  of  pumps.  The  first  class  is  often  pre- 
ferable for  intermittent  unwatering,  even  if  it  has  a  higher  operating 
cost,  for  where  the  existing  ore-hauling  and  hoisting  equipment  can  be 
utilized  to  move  the  water-buckets,  the  heavy  expense  of  installing 
pumps  is  obviated.  Where  air  compressors  are  already  installed,  the 
Pohle  air-lift  system  can  cheaply  be  applied  for  unwatering  the  upper 
levels. 

EDITOR'S  NOTE 

Chapters  VI  to  XVII  of  Mr.  Brinsmade's  treatise  "Mining  Without  Timber" 
have  been  omitted  from  this  Library  because  the  editor  has  not  felt  that  they  would 
be  of  sufficiently  definite  value  to  men  engaged  in  the  mining  of  coal.  The  com- 
plete work  is  a  standard  treatise  on  timberless  methods  of  mining  and  is  recom- 
mended to  all  who  wish  to  carry  their  studies  further. 


CHAPTER  XVHI 
PRINCIPLES  OF  MINING  SEAMS 

(a)  COMPARISON  OP  LONGWALL  AND  PILLAR  SYSTEMS 

In  the  mining  of  seams  of  coal  or  of  other  bedded  deposits  of  similar 
regular  thickness  over  large  areas  like  iron  ore,  gypsum,  salt,  etc.,  the 
two  systems  largely  used  are  "longwall"  and  "room  and  pillar"  or 
simply  "pillar."  The  longwall  system  extracts  all  the  coal  in  the  first 
operation  along  a  long  stretch  of  "wall"  or  face  and  allows  the  roof  to 
settle  gradually  behind  the  miners  upon  a  "  gob  "  or  partial  waste  filling. 
The  only  excavation  kept  open  besides  a  narrow  passage  at  the  working 
face  are  a  few  roads,  to  the  bottom  of  the  shaft  or  other  exit,  for  venti- 
lation and  for  the  tramming  of  coal  and  supplies.  In  the  pillar  system, 
on  the  contrary,  the  first  operation  consists  of  driving  roadways  to  ex- 
tract only  part  of  the  seam  in  rooms,  while  leaving  the  balance  in  the 
form  of  pillars  to  sustain  the  roof.  Later  the  pillars  are  drawn  or 
"  robbed  "  so  as  to  finally  recover  as  much  of  the  seam  as  possible. 

Either  system  may  be  pursued  "advancing"  or  "  retreating."  If 
advancing,  the  attack  of  the  longwall  or  the  robbing  of  the  pillar  system 
begins  next  the  safety  pillar,  left  to  protect  the  shaft  or  other  entrance, 
and  advances  outwardly  toward  the  boundry;  while  if  retreating,  the 
roadways  of  the  longwall  or  the  roads  and  rooms  of  the  pillar  system 
are  driven  to  the  outer  boundary  of  the  mine  before  the  "attack  of  face" 
or  the  "  robbing  of  pillars, "  respectively,  is  begun. 

The  two  broad  divisions  of  the  longwall  system  are  "continuous- 
face,"  in  which  the  face  is  kept  in  the  form  of  a  circle  or  similar  closed 
figure,  and  "  panel "  where  the  face  is  handled  in  panels  or  blocks,  along 
a  sufficient  stretch  for  free  roof  subsidence,  without  forming  a  closed 
figure.  These  divisions  have  each  several  varieties  and  often  shade  into 
each  other. 

The  pillar  system  has  three  varieties:  "room  and  pillar,"  where  the 
rooms  are  wider  than  the  pillars;  "stall  and  pillar,"  where  the  stall  or 
room  is  narrower  than  the  pillar;  and  "panel,"  where  the  mine  is  divided 
into  sections  or  panels,  separated  from  each  other  by  peripheral  pillars, 
and  each  is  divided  into  a  number  of  rooms  with  corresponding  pillars. 
In  some  mines  it  has  been  found  advantageous  to  combine  the  longwall 
and  pillar  systems  or  even  to  operate  them  separately  in  different  portions 
of  the  property. 

The  longwall  system  is  adapted  only  to  uniform  seams  with  roofs  of 

62 


PRINCIPLES    OF    MINING    SEAMS  63 

an  elastic  material  like  shale  or  sandstone  rather  than  those  with  a  blocky 
fracture  like  limestone.  Hitherto,  longwall  has  been  most  used  for 
working  seams  of  coal,  but  it  is  likely  hereafter  to  be  widely  applied  to 
other  deposits  like  the  Appalachian  iron  beds,  the  Michigan  copper 
amygdaloids,  or  the  Transvaal  gold  banket,  in  order  to  overcome  the 
obstacles  incident  to  great  depth.  In  coal  mining,  longwall  has  a 
particular  advantage  in  thin  seams  over  the  pillar  system,  because  the 
robbing  of  pillars  in  such  seams  in  usually  unprofitable,  and  longwall 
reduces  to  a  minimum  the  expense  of  driving  and  maintaining  the  road- 
ways. Longwall  also  gains  in  desirability  with  increasing  depth  where 
the  pillars  of  the  rival  system  must  continually  widen  and  thus  proportion 
ately  be  dearer  to  recover. 

Longwall  is  adapted  to  beds  containing  considerable  waste,  for  the 
waste  can  all  be  stored  underground  and  if  suitable  for  pack  walls  will 
obviate  the  use  of  timber  cogs.  In  Europe,  with  plenty  of  waste  for 
gobs  and  packs,  seams  as  thick  as  10  ft.  have  been  worked  in  one  slice 
by  longwall.  Less  timber  is  consumed  in  the  longwall  than  in  the  pillar 
system  because  in  the  latter  the  props  can  seldom  be  used  again.  The 
subsidence  of  the  longwall  roof  is  gradual  so  that  it  does  not  inflict  such 
breaks  in  the  formation,  to  let  in  water  or  to  damage  surface  structure, 
as  ensue  from  pillar-robbing.  By  longwall,  a  mine  can  be  developed 
more  quickly  and  more  cheaply,  and  more  lump  coal  and  a  higher  per- 
centage of  the  seam  can  be  excavated  in  less  time  than  by  pillar  work. 
In  longwall,  the  ventilation  system  is  cheaper  to  construct  and  to  main- 
tain, for  the  mine's  resistance  is  less;  few  or  no  explosives  are  needed; 
and  there  is  less  danger  from  falls  of  the  roof.  Longwall  requires  better 
trained  miners  than  pillar  work  but  a  miner's  output  is  greater.  After 
a  strike  of  nearly  six  months  recently  in  a  Western  coal  district, 
it  cost  the  pillar  mines  nine  times  as  much  to  clean  up  and  to  get  started 
again  as  it  did  similar  mines  where  the  longwall  system  prevailed.  This 
result  was  strictly  opposite  to  the  opinion  previously  held  by  many  on 
the  subject.  Longwall,  however,  is  unsuited  to  fluctuating  outputs,  for 
the  roof,  when  being  moved  at  all,  should  subside  uniformly  along  the 
face.  For  coal  mining,  longwall  is  gradually  superseding  pillar  work 
in  Europe  wherever  conditions  are  suitable.  America  is  bound  to  follow 
suit  as  soon  as  her  mines  become  deeper. 

(6)  COMPARISON  OF  THE  RETREATING  AND  ADVANCING  SYSTEMS 

The  retreating  system  of  mining  seams  is  rapidly  supplanting  its 
rival,  the  advancing  system,  in  European  mines,  but  in  America  the 
author  knows  of  no  case  of  its  use  in  longwall  and  of  comparatively  few 
cases  in  pillar  working.  On  inquiring  why  the  advancing  system  is  still 
preferred,  the  only  two  reasons  to  be  found  for  its  use  in  opening  a  mine 


64  MINING    WITHOUT    TIMBER 

are  that  it  requires  less  capital  and  less  time.  Everything  else  is  against 
the  advancing  system  which  has  a  smaller  percentage  of  mineral  recovery, 
a  worse  control  of  roof,  ventilation,  and  drainage,  a  greater  liability  to 
gas  and  dust  explosions,  a  higher  cost  of  maintenance  of  roadways,  a 
larger  timber  consumption,  and  a  smaller  output  from  an  equal  developed 
area  at  the  face. 

The  mistake  of  sacrificing  safety,  mineral  and  profit  per  acre  in  order 
to  get  an  output  quickly  at  minimum  cost  may  be  unavoidable  for  small 
weak  enterprises,  but  no  valid  excuses  can  be  made  by  strong  companies 
which  continue  to  persue  such  a  penny-wise,  pound-foolish  policy.  The 
driving  of  entries  to  the  boundary  to  inaugurate  the  retreating  system 
for  either  longwall  or  pillar  working  completely  explores  the  traversed 
territory.  These  entries  expose  the  seam's  faults,  rolls  and  irregu- 
larities, and  thus  indicate  both  the  lowest  points  of  the  floor  for  the  loca- 
tion of  sumps  and  pumps,  and  the  high  points  of  the  roof  where  may  be 
placed  churn-drill  holes  for  the  escape  of  gas  or  safety  shafts  with  ladders 
to  serve  as  natural  ventilators  when  the  fan  is  idle. 

With  the  retreating  system  not  only  are  there  no  old  gobbed  areas 
within  the  active  workings  to  generate  foul  gases  and  fires,  but  before 
stoping  begins  and  fills  up  the  mine  with  men,  the  seam  has  been  per- 
forated everywhere  by  the  entries,  and  most  of  its  water-channels  and 
pockets  or  feeders  of  gas  have  been  discovered  and  placed  under  control. 
In  retreating,  when  stoping  begins,  drainage,  ventilation  and  tramming 
are  covering  the  whole  area  of  the  property,  and  are  at  a  maximum; 
and  all,  especially  the  two  latter,  tend  to  grow  less  as  the  working  area  is 
contracted,  while  the  advancing  system  implies  a  continual  extension  of 
the  area  covered  by  each.  The  maintenance  of  entries  is  a  serious  ex- 
pense in  an  advancing  system,  as  they  must  not  only  be  constantly  re- 
brushed,  but  are  liable  to  develop  irregularities  from  squeeze  which 
make  uniform  tramming  grades  difficult  to  maintain.  The  final  capital 
cost  of  entries  is  the  same  in  either  system  but  the  cost  of  maintenance 
for  retreating  is  only  a  fraction  of  that  for  advancing.  This  gain  alone 
will  often  more  than  offset  the  earlier  outgo  of  capital  requisite  for  the 
former  system. 

With  the  advancing  system,  coal  is  apt  to  be  lost  even  by  longwall, 
while  the  history  of  even  recent  pillar  working  in  America  indicates  that 
an  average  of  hardly  70  per  cent,  of  the  seam  is  recovered.  To  obviate 
the  only  two  drawbacks  to  retreating,  the  need  of  much  capital  and  time, 
the  two  systems  can,  in  pillar  working,  be  easily  combined  temporarily, 
by  opening  off  enough  rooms  from  the  advancing  entries  to  maintain  a 
modest  output  until  the  boundary  is  reached,  where  pillar-drawing  can 
then  be  started,  and  the  regular  retreating  system  inaugurated.  In 
longwall  working,  the  combination  of  advancing  and  retreating  is  less 
simple  because  of  the  complications  it  is  liable  to  cause  in  the  control  of 


PRINCIPLES    OF    MINING    SEAMS  65 

roof,  especially  with  the  continuous  face  method;  but  with  the  panel 
layout,  it  can  be  effected  in  those  cases  where  the  roof  is  flexible  enough 
to  permit  the  longwall  operation  of  isolated  panels  of  moderate  size. 

The  longwall  practice  cited  in  the  next  chapter  is  all  on  the  advancing 
system  owing  to  the  lack  of  retreating  examples  in  America,  but  the 
advance  layouts  described  can  readily  be  transformed  into  those  for 
retreat  by  merely  starting  the  initial  longwall  face  at  the  boundary  of 
the  property  instead  of  at  its  entrance. 

(c)  MINING  BY  ROOF-PRESSURE 

Blasting  must  ever  be  a  danger  in  a  colliery;  and  all  practical  sub- 
stitutes not  involving  the  creation  of  flame  or  high-temperature  gases  are 
to  be  welcomed.  The  different  forms  of  wedges,  hydraulic  cartridges, 
lime  cartridges,  and  other  appliances  of  like  purpose  have  all  received 
full  attention,  but  little  has  been  written  on  Nature's  own  solution  of 
the  difficulty;  viz.,  roof  pressure,  and  its  systematic  and  scientific 
utilization. 

Any  bed  or  seam  is  subjected  to  a  certain  compressive  force  owing  to 
the  weight  of  the  superincumbent  strata:  if  a  portion  of  the  bed  is  removed 
and  no  artificial  means  of  supporting  the  excavation  attempted,  a  "  center 
of  relief"  is  established,  the  roof  and  floor  of  the  cavity  move  together, 
and  the  coal  (or  other  material)  round  and  about  the  cavity  is  cracked 
and  crushed  by  the  roof  weight,  and  eventually  some  of  it  forced  out  into- 
the  open  space.  If  the  coal  surrounding  the  excavation  had  been  under- 
cut, it  would  have  fallen  under  the  action  of  the  roof  weight  sooner  and 
in  better  condition;  but,  had  the  undercut  been  too  deep,  the  coal  would 
have  fallen  en  masse,  and  have  necessitated  manual  labor  in  breaking  to 
a  size  suitable  for  removal. 

The  cleavage  of  the  coal  must  be  studied  in  order  to  determine 
its  behavior  under  roof  pressure.  The  terms  bord  (or  face)  and  end 
(or  butt)  are  pretty  universally  employed  in  application  to  a  coal  face 
advancing  with  its  length  parallel  to  the  planes  of  main  cleavage  or  cleat, 
and  perpendicular  to  those  planes  respectively.  It  is  well  known  to 
every  collier  that  bordways  is  the  easiest  direction  of  advance;  but  coal 
so  hewn  is  most  likely  to  result  in  a  high  proportion  of  slack.  On  the 
other  hand,  the  coal  is  strongest  end-on;  is  hardest  to  hew,  but  is  most 
likely  to  result  in  a  large  percentage  of  lump  when  so  obtained. 

The  mode  of  fracture  of  the  roof  also  needs  attention.  The  forces 
which  induced  the  cleat  into  the  coal  had,  in  the  generality  of  cases,  a 
similar  effect  on  the  strata  above,  causing  an  incipient  cleavage  in  it 
coincident  in  direction  with  the  cleat  of  the  coal.  For  this  reason,  the 
maintenance  of  a  long  straight  face  absolutely  bord  is  almost  an  impos- 
sibility; at  such  a  face  the  roof  would  be  beyond  control,  would  break 


66  MINING    WITHOUT    TIMBER 

off  "short"  against  the  face  (Fig.  108),  and  would  not  only  be  a  constant 
source'  of  danger,  but  would  take  most  of  the  useful  weight  from  the  face. 
This  last  fact  was  recognized  very  early  by  coal  miners  and  wishing 
to  combine  the  easy  bord  direction  of  advance  with  a  better  control  of 
roof,  they  instituted  the  stepped  face  (Fig.  119).  Since  the  mean  line 
of  advance  in  the  case  shown  in  the  figure  runs  some  30  deg.  from  bord, 
it  follows  that  the  roof  will  break  parallel  to  this  line.  Stepped  long- 
wall  has  many  disadvantages;  first  among  which  must  be  placed  the 
fact  that  the  stepped  face  is  unsuited  to  machine  holing.  Secondly,  out- 
standing points  of  coal,  such  as  K,  Fig.  119,  receive  an  undue  roof  weight 
and  become  crushed.  While  at  the  other  extreme  we  have  points  such 


Fid.  108. — Effect  of  pressure  on  roof. 

as  ra,  Fig.  119,  too  far  back  and  too  well  protected  for  the  roof  weight 
to  act  usefully  there,  where  also  the  coal  is  bound  on  two  sides  (along 
the  face  and  down  the  step),  and  correspondingly  hard  to  hew.  Again, 
since  the  packs  have  to  be  built  close  against  the  side  of  the  step  to 
support  the  coal  (the  space  between  the  two  is  seldom  more  than  2  ft., 
often  less)  'the  ventilating  current  suffers  from  such  restrictions — an 
effect  which  is  further  augmented  by  a  frictional  loss  brought  in  by 
the  air  being  forced  to  travel  a  zig-zag  path.  Stepped  longwall  is 
giving  way  in  places  to  the  straight  "half-on"  face,  which  allows  of 
machine  holing  and  a  well-controlled  roof. 

Further  factors  which  must  receive  consideration  in  a  discussion  of 
the  effects  of  roof  pressure,  beyond  those  outlined  above,  are : 

1.  The  nature  of  the  seam. 

2.  The  nature  of  the  floor  and  roof. 

3.  The  rate  of  advance  of  the  face. 

4.  The  amount  of  dip  of  the  strata  and  the  direction  of  dip  as  com- 
pared with  the  direction  of  the  cleat. 

To  utilize  the  roof  weight  to  the  best  advantage,  the  coal  must  be 
undercut  to  a  certain  uniform  depth,  such  that  when  the  sprags  are 
withdrawn  the  coal  falls  with  a  vertical  fracture  from  the  back  of  the 
undercut,  without  any  extraneous  aid  by  blasting,  or  even  wedging. 

To  achieve  this,  the  undercut  must  generally  be  deeper  than  what  is 


PRINCIPLES    OF    MINING    SEAMS  67 

considered  advisable  by  hand;  hence,  we  must  depend  on  the  machine 
to  make  this  desideratum  an  actuality.  At  the  Altofts  Colliery,  Nor- 
manton,  Yorkshire,  they  have  succeeded  in  almost  dispensing  with  blast- 
ing by  holing  5  ft.  6  in.  under  in  a  flat  3  ft.  3  in.  seam,  1500  ft.  below  the 
surface.  With  a  hard  seam  a  much  deeper  undercut  than  5  1/2  ft. — 
perhaps  7  1/2  ft.  or  8  ft.  in  some  cases — would  be  found  necessary  if 
blasting  were  to  be  abolished;  such  a  depth  would  if  it  became  anything 
like  general,  cause  the  abandonment  of  disk  machines  in  favor  of  those 
of  either  the  bar  or  "puncher"  type. 

A  most  important  advantage  of  the  coal-cutting  machine  lies  in  the 
straightness  and  length  of  face  necessary  for  successful  application:  the 
straightness  of  the  face  enables  the  timbering  to  be  absolutely  systematic; 
and  this  factor  together  with  the  great  length  of  the  face  allows  of  the 
roof  pressure  to  be  controlled  and  utilized  with  precision.  Where  faults 
are  absent,  the  longer  the  longwall  face,  the  more  effectively  may  the 
roof  pressure  be  employed  as  the  means  of  breaking  down  the  coal. 

The  weight  of  the  roof  is  not  the  only  force  in  action  on  the  coal; 
before  the  coal  is  worked  the  pressure  of  the  floor  is  exactly  equal  and 
opposite  to  that  of  the  roof  on  the  seam;  when  an  excavation  is  made  in 
the  seam,  the  floor,  expanding  on  being  relieved  of  much  of  its  com- 
pression, exerts  an  upward  force  which  at  first  is  of  the  same  intensity 
as  the  roof  pressure  but  which  becomes  dissipated  sooner  than  the 
latter;  nevertheless  the  floor  pressure  is  often  of  service  to  the  miner 
and  is  taken  advantage  of  at  many  collieries  where,  owing  to  there  being 
a  suitable  band  of  dirt  at  or  near  the  top  of  the  seam,  overcutting  is 
resorted  to  in  lieu  of  undercutting.  Coal  so  obtained  often  is  in  better 
condition  than  coal  obtained  by  undercutting. 

A  treacherous  roof,  which  breaks  and  falls  immediately  the  weight 
comes  on  it,  rendering  timber  of  little  avail,  is  an  undoubted  evil.  Much 
may  be  done  in  the  way  of  palliation,  however,  by  quickening  the  rate  of 
advance  of  the  face,  and  proportionally  shortening  it  to  maintain  a 
uniform  output  (the  same  is  also  advisable  in  the  case  of  a  soft  seam).  It 
has  often  been  found  effective  in  keeping  up  a  bad  roof  to  leave  a  thin 
strip  of  coal  against  the  roof.  The  device  seems  to  act  something  like  a 
plaster  on  a  wound:  it  has  an  effect  out  of  all  proportion  to  the  slight 
increment  of  strength  it  supplies :  its  action  is  to  prevent  the  slacking  and 
slipping  of  the  roof;  to  maintain  it  in  its  entirety. 

Just  previous  to  installing  machine  cutting  into  a  colliery,  experi- 
ments must  be  made  to  ascertain  the  depth  of  holing,  such  that  when  the 
sprags  are  withdrawn  the  roof  pressure,  aided  by  the  weight  of  the  coal 
undercut,  is  sufficient  to  break  off  the  coal  at  the  back  of  the  holing.  The 
result  arrived  at  will  be  somewhat  (6  in.  or  a  foot)  short  of  the  correct 
figure,  inasmuch  as  the  coal  face,  when  undercut  by  machine,  will 
advance  two  or  three  times  as  speedily  as  when  the  holing  is  done  by 


MINING    WITHOUT    TIMBER 


hand,  and  hence  the  roof,  not  having  the  time  to  weigh  so  heavily  will 
require  a  larger  surface  on  which  to  act. 

The  coal  when  the  sprags  or  wedges  are  removed  must  fall  not  in  a 
solid  block  (Fig.  109),  but  well  cleaved  (Fig.  110)  and  ready  for  immediate 


FIG.  109. — Undercut  coal,  fallen  en  masse. 


filling.  Should  the  coal  fall  as  exemplified  by  Fig.  109,  the  defect  can 
generally  be  remedied  by  turning  the  face  more  toward  bord,  or  by 
lessening  the  rate  of  advance  (the  former  method  in  preference),  and 
experiment  in  that  direction  should  be  made  at  once. 


FIG.  110.— Undercut  coal,  fallen  in  blocks. 

« 

Judging  from  present-day  experience,  little  need  be  feared  on  the 
count  of  coal  cutting  when  the  depths  of  our  mines  become  excessive; 
indeed,  under  the  heavy  roof  pressures  then  in  action  the  use  of  explo- 
sives at  the  coal  face  is  likely  to  be  abolished,  and  the  depth  of  holing 


PRINCIPLES  OF  MINING  SEAMS  69 

necessary  will,  if  anything,  be  less  than  that  at  present  in  vogue.  In 
very  deep  mines,  however,  it  has  been  found  that,  although  the  character 
of  the  seam  remains  the  same,  the  percentage  of  slack  increases  with  the 
depth.  Before  the  Royal  Commission  on  Coal  Supplies,  Mr.  Martin 
opined  that  an  increase  of  depth  from  1200  to  2400  ft.  would  result  in 
an  increase  in  the  percentage  of  slack  from  the  same  seam  of  5  per  cent. ; 
and  at  Pendleton  Colliery,  while  the  coal  was  worked  at  depths  of  less 
than  2  500  ft.,  the  percentage  of  slack  was  21.5,  but  when  the  workings 
had  reached  the  depth  of  between  3000  and  3500  ft.,  the  proportion  had 
increased  to  39  per  cent.  This  is  merely  another  way  of  stating  that  the 


FIG.   111. — Roof  pressure  when  mining  to  rise. 

roof  pressure  has  been  too  severe  for  the  coal,  and  to  mitigate  such  an 
effect  the  coal  should  be  got,  as  far  as  possible,  by  machine,  working  end- 
on,  and  with  a  rapid  rate  of  advance. 

The  effects  of  dip  on  the  action  of  the  roof  pressure  is  important. 
In  a  working  proceeding  full  rise,  experience  tells  us  that,  other  things 
being  equal: 

1.  Hewing  is  easier. 

2.  Work  is  more  dangerous  (from  falls  of  roof  and  face). 

3.  More  slack  is  produced  than  in  a  similar  working  in  a  flat  seam. 
Hewing  is  easier  for  the  reason  that  both  the  roof  pressure  and  the 

weight  of  the  coal  have  more  total  useful  effect  in  a  rise  working  than  in 
any  other.  In  Fig.  Ill,  by  means  of  the  undercut  A  B  a.  wedge-shaped 
block  of  coal  A  B  C  D  is  undermined,  if  sprags  or  wedges  be  placed 
under  the  mouth  of  the  undercut,  the  triangular  block  A  D  E  is  still 
unsupported,  giving  us  at  once  the  reason  for  the  liability  to  falls  of  face 
in  such  a  working,  and  also  demonstrating  the  need  for  the  cocker  sprag 
(shown)  or  equivalent  means  of  supporting  the  face.  The  action  of  the 
roof  is  two  fold.  There  is  a  pressure  P,  acting  normally  to  the  plane 


70  MINING    WITHOUT   TIMBER 

of  the  seam;  there  is  a  thrust  T,  acting  in  the  direction  of  dip,  tending 
to  make  the  roof  slide  over  the  face  toward  the  empty  space  behind  it. 

The  force  T  is  evidenced  in  the  fact  that  a  fracture  in  the  roof  of  a 
rise  working  "  gapes,"  owing  to  the  lower  side  having  moved  slightly  down 
under  the  influence  of  T.  Thus  it  is  that  falls  of  roof  are  more  prevalent 
in  rise  workings  than  in  any  other;  the  side  thrust  T,  not  only  quickly 
breaking  up  the  roof,  but  also  widening  the  joints  the  better  to  allow 
severed  slabs  to  fall. 

It  is  largely  to  this  side  thrust  that  we  owe  the  production  of  slack 
which  is  one  of  the  disadvantages  of  rise  working;  grinding  is  introduced, 
a  far  more  effective  slack  producer  than  mere  normal  pressure. 

The  resultant  action  of  the  forces  P  and  T  on  the  coal  may  be  best 
represented  by  the  single  force  R.  The  direction  of  R  cannot  be  accu- 
rately assigned,  but  it  lies  somewhere  between  the  normal  (P)  and  the 
perpendicular  (shown  dotted),  and  its  position  is  probably  somewhere 
as  shown;  its  magnitude,  by  the  parallelogram  of  forces  is  simply\/P2  +  T2. 
Considering  the  coal  acted  on  by  R  aided  by  the  weight  of  the  coal  itself, 
no  further  demonstration  is  needed  of  the  reason  of  the  ease  experienced 
in  working  coal  to  the  rise 


FIQ.  112. — Roof  pressure  when  mining  to  dip. 

Rise  working  would  be  rendered  safer,  and  less  slack  would  be  produced 
if  a  rapid  rate  of  advance  were  maintained,  and  to  compensate  for  the 
lessening  of  roof  pressure  which  would  result,  a  deeper  undercut  would 
be  necessitated.  Carefully  built  pack  walls  are  also  highly  advisable  in 
a  seam  liable  to  produce  slack,  and  thus  especially  in  rise  workings. 

In  the  dip  working  (Fig.  112),  the  action  of  T,  the  side  thrust,  is  much 
less  important;  the  tendency  is  there,  but  the  action,  so  far  as  the  grinding 
of  the  coal  is  concerned,  is  nil.  Also,  any  line  of  break  appearing  in  the 
roof  is  closed,  instead  of  opened,  by  the  slight  lateral  movement  of  the 
roof  over  the  gob  or  goaf:  hence,  working  to  the  dip  of  the  seam  is, 
generally  speaking,  the  safest  of  all  directions  of  working  the  coal.  As 
before  the  resultant  roof  pressure  acts  slightly  down-hill  (shown  at  R), 
causing  the  coal  in  this  case  to  be  difficult  to  hew. 

The  coincidence  or  noncoincidence  of  the  direction  of  cleavage  and 
the  direction  of  dip  is  a  factor  of  importance,  influencing  the  behavior  of 


PRINCIPLES    OF    MINING    SEAMS  71 

the  coal  under  the  roof  pressures.  If  the  directions  bord  and  dip  coincide 
a  rise  working  is  doubly  easy,  but  the  face  will  need  stepping  or  the  roof 
will  be  beyond  control;  also,  under  these  conditions,  dip  working  will  be 
facilitated,  and  generally  the  face  in  such  a  working  may  be  maintained 
straight,  owing  to  the  side  thrust  closing  the  jointings  in  the  roof.  On 
the  other  hand,  should  the  directions  end  and  dip  coincide,  the  easiest 
mode  of  advance  will  not  be  full  rise  but  in  some  direction  between  that 
line  and  the  strike  of  the  seam,  while  the  difficulty  of  working  directly 
to  the  dip  will  be  intensified.  Intermediate  between  these  extremes 
there  is  an  infinite  number  of  angles  at  which  the  directions  of  dip  and 
cleat  may  lie,  every  case  needing  special  consideration  and  experiment. 


CHAPTER  XIX 
ADVANCING  LONGWALL  SYSTEMS  FOR  SEAMS 

EXAMPLE  49. — SPRING  VALLEY  BITUMINOUS  COLLIERIES,  BUREAU 
COUNTY,  ILLINOIS 

(See  also  Example  5.) 

Thin  Flat  Seam  at  500-/i.  Depth.  Advancing  with  Continuous  Face  by 
Scotch  System.  Loading  into  Cars. — The  Illinois  coal  reports  show  that 
over  5,000,000  tons  are  produced  in  the  longwall  field  or  about  12  per  cent, 
of  the  States'  total  output.  Most  of  the  longwall  mining  is  done  in  the 
prairie-like  counties  of  Bureau,  Grundy,  and  La  Salle.  Here  the  coal  seams 
are  remarkable  not  only  for  their  variety  and  quality,  but  in  their  free- 
dom from  horse-back,  faults,  and  other  irregularities,  which  are  encount- 
ered elsewhere.  The  mines  are  developing  a  3  1/2-ft.  seam,  called  com- 
mercially "third-vein  coal,"  which  is  350  to  500  ft.  beneath  the  surface. 
Overlying  the  seam  is  a  flexible  shale  4  to  9  in.  thick  and  underlying  it  is 
6  to  24  in.  of  fireclay. 

The  Scotch  system  extracts  all  the  coal  in  the  first  operation,  com- 
mencing at  the  periphery  of  the  shaft  pillar  and  mining  out  the  whole 
seam  toward  the  property  limits.  In  Fig.  113,  the  ideal  plan  of  the  lay- 
out of  a  Spring  Valley  mine  350  ft.  deep,  h  is  the  hoisting,  a  the  air-shaft, 
and  km  the  shaft  pillar,  600  ft.  square,  over  which  are  located  the  shaft 
house,  shops,  and  other  necessary  surface  structures.  To  open  out  the 
seam  for  longwalling,  the  pillar  is  first  cross-cut  by  the  two  headings  gn 
and  pq,  the  former  being  double-tracked  to  permit  the  handling  of  cars  to 
and  from  shaft  h.  Roads  are  uniformly  9  ft.  wide,  excepting  at  the  part- 
ings where  they  are  14  ft.  wide  to  hold  two  tracks.  These  peripheral 
headings  are  driven  from  the  points  g,  n,  p,  and  q,  and  by  widening  them 
inbye,  the  first  longwall  face  is  begun.  At  first  the  face  is  not  continuous, 
but  in  arcs  like  those  dotted  around  points  g,  n,  p  and  q,  but  as  the  arcs 
move  inbye  they  finally  meet  and  form  one  continuous  circle.  Soon  the 
face  has  advanced  sufficiently  to  allow  the  first  break  to  be  made  in  the 
roof  around  the  shaft  pillar,  so  that  thereafter  advantage  can  be  taken  of 
the  pressure  from  the  descending  roof  to  mine  the  coal  as  in  Fig.  114. 

Fig.  1 13,  shows  the  longwall  face  after  it  has  advanced  some  distance 
from  the  shaft-pillars  and  appears  as  the  circle  1-4-7-10.  The  face  is 
reached  through  the  main  roadways  1,  2,  3,  etc.,  by  which  it  is  divided 
into  approximately  equal  spaces.  The  essential  feature  of  the  Scotch 

72 


ADVANCING    LONGWALL    SYSTEMS    FOR    SEAMS 


73 


system  by  which  it  differs  from  the  other  continuous-face  longwall  system, 
"  the  Rectangular,"  is  the  turning  off  of  cross-roads  at  45-deg.,  angles  from 
the  main  directions  for  roads,  g-n  and  p-q.  The  road  e-3  was  formerly 
the  extension  of  gk,  but  its  present  position  makes  possible  a  gentler 
curve  for  haulage  into  gh  and  permits  the  use  of  the  40-deg.  frog  of  the 
other  45-deg.  turnouts.  The  layout  of  roads  is  based  on  a  room  about 
42  ft.  wide  at  the  face  (as  at/)  which  corresponds  to  60  ft.  along  the  road. 


Overcast 
Door  or  Curtain 
Fio.   113. — Plan  of  Scotch  Longwall  system,  Illinois- 

c-4.  The  angle-roads  are  turned  off  at  225-ft.  intervals  from  a  cross-road 
as  e-3.  Experience  has  shown  that  a  room  can  not  exceed  a  length  of  225 
ft.  with  out  having  its  track  rebrushed  before  completion.  Hence,  the 
rooms  of  an  angle-road  like  cd  are  abandoned  as  soon  as  they  are  cut  off 
by  the  next  road  r-4.  Crossroads  e-3  and  y-5  are  1178  ft.  apart,  so  that 
angle-roads  c-d  and  d-g  can  each  be  840  ft.  long  to  contain  exactly  14 
rooms. 

The  direction  of  the  air  current  is  shown  by  arrows.  The  shaft  a  is  the 
downcast,  the  shaft  h  is  the  upcast;  roads  12,  2,  6,  and  8  are  intake,  and  the 
other  roads  are  return  airways  along  their  inbye  portions.  In  the  intake 


74 


MINING    WITHOUT    TIMBER 


roads,  double  doors  are  placed  at  t,  t',  y,  and  y'  while  single  doors  are  placed 
at  points  with  strong  draft  like  z,  z' ,  etcv  to  hold  the  air  along  the  working 
face.  Fire-proof  burlap  curtains  are  used  in  rooms  where  the  draft  is  not 
strong  enough  to  force  them  up.  The  ventilation  is  excellent  and  easily 
regulated.  Trap  boys  are  stationed  at  the  double  doors  and  also  at  cross- 
roads where  collisions  are  liable  to  occur  from  trains  approaching  in  oppo- 
site directions. 


FIG.  114.— Roof  sinking  behind  Longwall  face. 

For  haulage  the  grade  is  nearly  level;  big  mules  are  used  on  the  main 
roads  and  small  ones  for  gathering.  At  No.  4  mine,  7  gathering  mules 
with  two  cars  apiece  haul  250  cars  daily  from  the  face  to  the  second  parting, 
thence  4  miiles  haul  trains  of  30  cars  to  the  first  parting  whence  they  are 
hauled  in  similar  4-mule  trains  to  the  shaft  bottom.  The  total  output  of 
1000  tons  is  hauled  in  the  day  shift  by  about  40  mules,  the  face  being 
about  a  mile  distant  from  the  shaft.  The  cars  are  of  wood,  weigh  1100 
lb.,  hold  2700  lb.,  of  coal  and  run  on  a  track  of  42-in.  gauge,  laid  with  16- 
Ib.  rails  except  at  the  bottom  of  the  shaft. 


FIG.  115. — Plan  of  packs  and  tracks  at  Longwall  face. 


Fig.  115  is  a  plan  at  the  face  showing  two  room  "gateways,"  or 
room  roads  with  tracks  turned  off  at  45  deg.  from  the  "  Timbered  Branch 
Road."  Halfway  between  the  gateways  is  the  "  mark"  a  which  separates 
a  room  into  two  21-ft.  halves,  each  assigned  to  one  miner  who  both  mines 
and  loads  his  coal  into  a  car  on  the  nearest  track  at  c.  The  undercut  is 
made  by  hand  pick,  from  a  crouching  position,  in  the  floor;  when  the  lat- 
ter is  of  sandstone,  the  coal  itself  must  be  grooved.  ,  If  the  roof  is  working 


ADVANCING    LONGWALL    SYSTEMS    FOR    SEAMS 


75 


properly,  which  is  ascertained  by  sounding  the  face  with  a  hammer,  the 
undercut  need  only  be  20  in.  deep.  Otherwise  a  depth  of  5  ft.  is  some- 
times necessary.  The  clay  cuttings  are  thrown  back  into  the  gob. 

The  packwalls  along  the  gateways  are  6  ft.  wide  and  do  not  approach 
nearer  than  2  ft.  to  the  face  in  order  not  to  obstruct  the  air-current. 
They  are  built  of  slate  brushed  from  the  roof  (see  Fig.  109),  by  a  hand 
pick  to  a  height  of  about  6  1/2  ft.  above  the  tracks.  The  roof  of  the 
haulage  roads  sinks  as  the  face  advances  and  must  be  continually  re- 
brushed  for  slate,  which  is  partly  stored  in  an  abandoned  room  and  partly 
hoisted  in  cars  to  the  extent  of  10  per  cent,  of  the  coal-car  hoist.  The 
rebrushing  of  the  roads,  the  repair,  of  the  track  and  the  retimbering,  is 
done  by  company  men  on  day's  pay,  but  initially  the  miners  brush  the 
gateways,  lay  the  tracks,  build  up  the  packwalls,  and  set  the  props  in  their 
own  rooms  as  part  of  their  contract  price  per  ton  of  coal  loaded. 

The  undercut  is  held  up  by  sprags  (see  Fig.  116)  until  the  whole  42-ft. 
ace  of  a  room  has  been  completed.  When  the  sprags  are  knocked  out 


Section  along 


FIG.  116. — Section  of  Longwall  face,  gateway  and  branch  road. 

and  the  undercut  coal  does  not  fall,  it  is  wedged  down  in  two  layers  by 
steel  wedges  starting  from  a  shear  made  at  the  breast  center  or  "mark." 
To  reach  the  car  from  the  mark,  the  coal  must  be  reshovelled  twice  in  the 
narrow  alley  between  props  and  face.  Lines  of  8-in.  by  4  1/2-ft.  props, 
about  3  ft.  apart,  are  set  along  the  face  (see  Fig.  114)  at  necessary  inter- 
vals and  some  of  these  are  not  recovered.  The  haulageways  are  timbered 
with  three-quarter  sets  which  in  certain  places  are  seen  to  be  cribbed  10  ft. 
high  above  the  caps  to  catch  up  roof-caves  that  broke  down  the  original 
timbering.  Steel  I  beams  are  used  for  caps  over  some  of  the  double 
track  partings  and  the  shaft  bottom  road  is  walled  with  masonry. 
Wooden  cogs  are  placed  at  acute  road  corners  as  c  (Fig.  113)  and  also  to 
replace  props  along  the  coal  face  where  the  roof  is  unusually  weak. 

Work  can  cease  on  part  of  the  coal  face  without  injury  to  the  balance 
if  care  be  taken  to  keep  the  whole  face  regular  and  convexly  curved,  for 
trouble  ensues  if  corners  or  concavities  are  allowed  to  develop.  A  fall 
or  creep  of  roof  at  the  face,  which  closes  up  the  space  between  gob  and 
coal,  may  occur  from  uneven  advances,  from  failure  to  build  up  the  pack 


76  MIXIN'G    WITHOUT    TIMBER 

walls,  from  unusually  weak  spots  in  the  roof,  or  from  long  shutdowns. 
The  face  is  reopened  as  "yardage"  work  by  driving  the  gateway  into  the 
coal  and  turning  off  a  heading  along  the  face  with  a  2-ft.  rib  between  it 
and  the  old  gob. 

The  haulageways  are  simply  gateways  selected  because  they  occur  at 
the  proper  intervals  of  the  layout  of  Fig.  113.  Near  the  face,  rooms  are 
always  advancing  in  two  directions  which  intersect  each  other  at  a  45-deg. 
angle.  Thus  room  zf,  as  it  advances,  is  cutting  off  all  the  rooms  between 
r-4  and  zf  which  are  then  abandoned  and  their  occupants  assigned  new 
working  places.  It  is  always  arranged  to  give  an  experienced  miner  a 
21-ft.  face  to  work  for  himself,  but  occasionally  he  takes  a  green  hand  as 
helper  and  pupil.  There  is  little  gas,  though  two  fire-bosses  are  employed. 
The  mining  and  hoisting  is  all  done  on  day  shift.  The  output  of  coal  is 
about  21/2  tons  per  man  employed  above  and  below  ground. 

EXAMPLE  50. — MONTOUR  IRON  MINES,  DANVILLE,  PA. 

Thin,  Sloping,  Shallow  Beds.  Loading  into  Cars. — The  long-wall 
method  of  minin  g  when  introduced  into  these  mines  was  but  little 
'used  in  this  country,  and  seldom  in  beds  as  thick  as  these,  with  breasts 
frequently  4  to  5  ft.  high.  Figs.  117  and  118  show  the  general  method. 

Levels  are  driven  90  ft.  apart,  and  the  face  of  each  gangway  should 
be  kept  in  advance  of  all  higher  gangways,  so  that  of  the  gangways  C, 
E,  and  H,  for  instance,  the  face  of  C  should  be  the  farthest  from  the  slope 
or  the  mouth  of  the  drift;  but  in  fact  this  is  seldom  done  here,  as  it 
necessitates  the  outlay  of  large  capital  before  any  return  is  realized. 

The  gangways  are  driven  7  to  10  ft.  wide  and  5  1/2  to  7  ft.  high,  so 
a  man  or  mule  can  go  erect.  The  lowest  gangway  C,  Fig.  117,  is  known 
as  a  "fast-end"  gangway,  as  it  is  driven  entirely  in  the  solid,  while  E,  H, 
etc.,  are  "loose-end"  gangways,  with  but  one  side  in  the  solid  and  the 
other  side  formed  by  the  stowing.  The  face  of  the  gangway  C  should  be 
kept  far  enough  ahead  so  that  blasting  there  will  not  interfere  with  the 
workmen  in  the  breast  D,  and  the  same  consideration  should  determine 
the  distance  of  the  breast  F  from  the  face  of  E. 

To  facilitate  the  loading  of  the  ore  into  cars  from  chutes,  the  gang- 
ways are  so  driven  that  the  roof  of  the  bed  will  lie  at  the  top  of  the  upper 
rib,  Fig.  118,  and  to  secure  the  proper  gangway  height  the  bottom  rock 
must  be  taken  up.  In  doing  this  along  the  lower  gangway  C  a  drainage- 
ditch  is  left  upon  the  lower  side.  When  first  driven,  the  gangway  C  is 
timbered  upon  the  upper  side,  but,  as  settling  takes  place,  the  props  are 
usually  broken,  and  it  is  necessary  generally  to  renew  them  and  to  blow 
down  the  roof  of  the  gangway,  which  frequently  settles  sufficiently  to 
obstruct  the  haulage-way.  Often,  however,  the  stowing  becomes  so 
tightly  packed  in  settling  that  retimbering  is  unnecessary.  In  the  soft 


ADVANCING    LONGWALL    SYSTEMS    FOR    SEAMS 


77 


ore,  by  reason  of  the  creeping  of  the  bottom,  the  gangway-props  must 
sometimes  be  renewed  several  times.  Breasts  or  rooms  D  and  F  are 
turned  off  at  an  angle  of  35  to  45  deg.  with  the  direction  of  the  gangway, 
depending  on  the  dip  of  the  bed.  The  breasts  are  24  to  30  ft.  long,  and 
usually  there  are  five  breasts  in  a  tier  between  two  gangways.  The 
height  of  the  breast  varies,  with  the  nature  of  the  ore,  from  2  to  5  ft.  In 
the  hard  limestone-ore  are  three  streaks  of  ore  which  are  taken  out  if 
sufficiently  rich;  but  if  the  ore  is  lean  the  central  streak  alone  is  taken 
out,  with  just  enough  rock  to  allow  the  mine  to  work  his  breast.  The 
hard  limestone-ore  and  the  block-ore  have  to  be  blasted,  but  the  soft-ore 
is  scraped  out  in  the  form  of  mud. 


FlQS.   117  AND   118. 

FIG.  117. — Long  section  of  stope,  Montour  mine. 
FIG.  118. — Cross-section  of  stope,  Montour  mine. 

Each  breast  is  worked  by  a  miner  and  one  laborer;  or  two  miners  will 
combine  and  work  two  breasts;  and,  sometimes,  one  miner  and  two  or 
three  laborers  will  work  two  breasts.  The  miner  working  the  top  breast 
of  a  tier,  such  as  D4,  Fig.  117,  also  drives  the  gangway  E,  takes  up  the 
bottom  rock  to  give  sufficient  height  for  haulage,  piles  the  stowing  care- 
fully on  the  lower  side  of  the  gangway,  and  prepares  the  road-bed  for 
the  track-layers,  for  which  additional  work  he  is  paid  extra. 

A  ditch  is  not  left -along  the  loose-end  gangway,  as  the  water  should 
drain  through  the  stowing  to  the  fast-end  gangway  C.  The  gob  is 
thrown  loosely  between  the  breast  and  gangway  below,  excepting  along 
the  chutes  G,  where  it  is  carefully  piled  to  support  the  roof.  A  chute  or 
gateway  G,  2  1/3  ft.  wide,  is  left  for  each  breast,  down  which  the  ore  is 
thrown  to  the  gangway  below;  it  is  sometimes  lined  with  boards,  but 
generally  a  carefully-built  dry  wall  of  gob  suffices.  The  ore  from  each 


78  MINING    WITHOUT    TIMBER 

breast  is  carried  to  the  chute  by  hand,  and  drawn  out  from  its  bottom 
into  cars  as  desired. 

There  is  usually  a  platform  at  the  chute  bottom  to  facilitate  the  load- 
ing; in  the  soft  ore  it  is  placed  directly  above  the  car,  but  it  is  nearer  the 
bottom  in  the  hard  ore,  and  sometimes  the  ore  is  simply  allowed  to  pile 
up  along  the  gangway.  When  necessary  to  prevent  the  air-current 
drawing  up  a  chute,  a  canvas  curtain  is  hung  loosely  over  its  mouth;  but 
ordinarily  only  the  last  five  inside  chutes  are  kept  open,  the  others  being 
boarded  up  and  filled  with  gob  when  the  gangway  E  has  advanced 
enough  to  receive  the  ore  from  the  next  higher  tier  of  breasts.  The 
gangways  E,  H,  etc.,  are  connected  with  the  fast-end  gangway  C  by 
"pitching  gangways"  K  driven  through  the  gob,  and  back  of  these  last 
the  gangways  E  and  H  are  abandoned,  so  that  the  fast-end  gangway  C 
is  the  only  one  kept  open  through  its  entire  length.  The  gangways  in 
the  soft-ore  are  timbered  and  lagged  on  sides  and  top,  but  in  the  hard 
limestone-ore  and  in  the  block-ore  it  is  generally  necessary  to  timber  the 
sides  only,  as  the  roof  is  of  good  slate  or  sandstone. 

The  props  are  placed  2  to  6  ft.  apart,  depending  upon  the  nature  of  the 
roof.  In  the  hard  fossil-ore  and  in  the  block-ore  the  breasts  are  not 
timbered,  excepting  when  necessary  to  protect  the  chutes,  as  the  gob 
fills  up  the  space  and  supports  the  top.  In  the  soft  fossil-ore  small  props, 
3  to  5-in.  dia.,  are  used  to  keep  up  the  top,  as  the  gob  does  not  fill  more 
than  one-third  of  the  vacant  space.  Heavy  timbers  are  usually  placed 
along  all  chutes.  The  timber  is  furnished  by  the  company,  but  the 
miners  set  it,  both  in  the  breasts  and  along  the  gangway,  and  as  it  is  cut 
on  company  property  it  is  cheap. 

All  general  work,  such  as  track-laying,  the  clearing  away  of  "falls," 
etc.,  is  done  by  miners,  detailed  for  each  separate  piece  of  work,  instead 
of  by  laborers,  and  for  such  work  the  miners  are  paid  per  diem.  All 
drilling  is  done  by  hand,  and  the  ventilation  is  secured  by  natural  draft 
through  chimneys.  In  the  drifts,  the  cars  are  either  pushed  by  hand  or 
hauled  by  mules  to  the  mouth,  while  in  the  slopes  the  mine-cars  are 
hoisted  to  the  top  by  second-motion  engines. 

The  mine-cars  are  2  ft.  deep,  4  ft.  long,  and  3  1/2  ft.  wide,  and  hold 
about  one  ton  of  ore.  The  gauge  is  30  in.  and  the  wheels  are  14-in. 
dia.  and  loose. 

Cost  of  Mining. — For  breast-work,  miners  are  paid  by  the  ton  and 
for  gangway-work  they  are  paid  tonnage  and  yardage. 

Tonnage  payments  depend  upon: 

(1)  The  nature  of  the  ore. 

(2)  The  height  of  the  breast. 
Yardage  payments  depend  upon: 

(1)  The  nature  of  the  ore. 

(2)  The  kind  of  gangway. 


ADVANCING    LONGWALL    SYSTEMS    FOR   SEAMS  79 

Since  the  nature  of  the  ore  in  the  fossil-beds  and  the  height  of  breast 
vary  so  irregularly,  it  is  almost  impossible  to  give  exact  figures  so  that 
chiefly  ratios  will  be  given.  Upon  a  basis  of  $1  per  ton  for  mining 
block  ore,  the  following  are  the  prices  paid  during  the  past  15  years: 

Block  ore  per  ton $1 . 00 

Block  ore  per  yard,  fast-end  gangway 4 . 00 

Block  ore  per  yard,  loose-end  gangway 1 . 70 

Hard  fossil  ore  per  ton 0 . 95 

Hard  fossil  ore  per  yard,  fast-end  gangway 6 . 25 

Hard  fossil  ore  per  yard,  loose-end  gangway 2 . 40 

Unskilled  labor  per  day 0 . 73 

Soft  ore  costs  to  mine  one-third  to  one-half  the  above  hard  ore  prices. 

One  ton  per  day  for  each  man  working  in  a  breast  is  considered  an 
average  output  for  a  shift  of  10  hours.  The  miner  pays  his  laborer,  or 
laborers,  per  diem,  at  the  above  rate.  In  gangway-work  the  average 
rate  of  advance  was  15  ft.  per  month  for  loose-end  gangways  and  7  ft. 
per  month  for  fast-end  gangways.  Owing  to  the  many  conditions  affect- 
ing the  rate  of  advance  along  the  gangways,  it  was  necessary  to  employ 
a  system  of  "allowances"  in  payment  of  gangway-yardage  so  as  to 
equalize  as  nearly  as  possible  the  pay  of  gangway-miners. 

EXAMPLE    51. — BULL'S    HEAD    ANTHRACITE    COLLIERY, 
PROVIDENCE,  EASTERN  PA. 

(See  also  Examples  5  and  59.) 

Thin,  Sloping,  Shallow  Seam;  Panel  System;  Loading  into  Buggies 
on  Endless  Rope. — The  coal  property  is  about  1200  ft.  square,  and  a 
section  of  the  measures  just  above  and  below  the  seam  being  mined 
longwall  is  approximately  as  follows: 

20     ft.,  slate  and  soil; 
21/2  ft.,  fireclay; 

1  ft.,  bone; 

5  ft.,  coal  seam; 
40     ft.,  slate; 

2  ft.,  sandstone; 

6  in.,  slate; 

"30  -in."  coal  seam; 
18     in.,  hard  slate; 
9    ft.,  soft  shale; 
18     in.,  hard  slate; 

3  1/2  ft.,  coal,  4-ft.  seam; 
90     ft.,  sandstone  and  slatej 

8    ft.,  coal,  Diamond  seam. 


80 


MINING    WITHOUT   TIMBER 


Below  the  Diamond  seam  occur  the  Rock  seam,  the  Fourteen-foot, 
and  the  Clark,  all  of  which  and  also  including  the  Four-foot  and  Dia- 
mond seams  had  been  worked  out  by  room  and  pillar  prior  to  begin- 
ing  to  mine  the  Thirty-inch  seam  by  longwall. 

Consequently,  the  rock  above  and  below  the  Thirty-inch  seam  was 
cracked  and  in  many  cases  out  of  place,  the  cracks  often  extending  to  the 
surface.  In  consequence  the  footing  for  props  was  most  insecure,  and 
although  the  cover  above  the  Thirty-inch  seam  was  only  about  75  ft., 
it  was  impossible  to  hold  it  by  timbering  and  it  would  have  been  prob- 
ably impossible  to  take  out  the  coal  by  room  and  pillar.  The  longwall 
method  of  working  as  developed  by  Supt.  Vipond  is  shown  in  Fig.  119. 
A  rock  slope  was  driven  up  from  the  Four-foot  seam  at  a  slight  pitch  so 


FIG.  119. — Plan  of  Longwall  system,  Bull's  Head  colliery. 

that  empty  cars  can  be  hauled  up  the  pitch  by  mules.  From  the  head 
of  this  pitch  the  gangway  a  was  driven  31  ft.  wide  and  5  to  6  ft.  high, 
bottom  rock  being  taken  up  to  give  sufficient  height.  At  the  same 
time  the  parallel  airway  6  was  driven  and  ventilation  secured  by 
means  of  the  headings  shown  through  the  gangway  pillars.  The  airway 
b  connects  by  a  passageway  V  with  a  ventilation  shaft  from  the  under- 
lying Four-foot  seam.  The  rock  obtained  in  driving  the  airway  is 
piled  in  walls  along  both  sides  of  the  airway.  The  rock  resulting  from 
taking  up  the  bottom  during  the  driving  of  the  gangway  is  built  into 
a  continuous  wall  c  along  the  lower  side  of  the  gangway  and  into  walls 
d  16  ft.  wide  along  the  upper  side  of  the  gangway.  Through  the 
upper  walls  d  are  passageways  e  which  are  9  ft.  wide  and  are  spaced 
125  ft.  between  centers.  These  passageways,  called  gateways,  have 
loose  walls  b  8  ft.  wide  on  each  side,  thus  making  the  total  width  of  the 
gateway  and  the  walled  space  25  ft.  The  gateways  are  driven  the  same 
height  as  the  gangway  for  a  short  distance  in  from  the  gangway  so  as 
to  provide  a  place  for  the  mine  car  to  stand  while  it  is  being  loaded  and 


ADVANCING    LONGWALL    SYSTEMS    FOR    SEAMS 


81 


out  of  the  way  of  traffic  along  the  gangway.  This  distance  depends 
upon  conditions,  but  is  usually  not  over  40  ft.  Above  this  point  the 
gateway  is  made  only  the  height  of  the  coal  and  the  overlying  slate, 
that  is,  about  36  in.  The  coal  is  overlaid  by  about  6  in.  of  slate  which  is 
always  taken  down,  and  it  is  this  which  furnishes  the  greater  part  of  the 
material  needed  for  building  the  pack  walls  along  the  greater  length  of 
the  gateways  and  along  the  face  as  will  be  described  later. 

The  method  of  opening  out  a  face  is  shown  at  A.  Strips  are  taken 
off  the  face  parallel  to  the  gangway  and  the  gangway  walls  d,  and  as 
soon  as  sufficient  width  is  secured  between  the  gangway  wall  d  and  the 
face  of  the  coal,  a  track  h  is  laid  as  near  to  the  face  as  possible  so  that  it 
will  not  interfere  with  the  work  of  the  miners.  This  track  has  a  gauge 


Fia.  120. — Single  winch,  Bull's  Head  colliery. 


of  2  ft.  3  in.,  is  laid  with  25-lb.  rails,  which  are  10  ft.  long.  The  rail 
sections  are  joined  by  two  fish-plates,  one  placed  on  each  side  of  the 
flange.  The  rails  are  held  together  by  iron  bridles  which  are  laid  directly 
on  the  bottom.  On  this  track  is  a  small  buggy  into  which  the  coal  is 
shoveled.  This  buggy  is  moved  by  an  endless  wire-rope  operated  from 
a  point  i  on  the  gateway  as  follows:  A  cast-iron  wheel  a,  Fig.  120,  18  in. 
in  diameter  and  having  a  groove  2  in.  deep,  is  held  in  a  wooden  frame  b. 
At  the  bottom  is  a  pointed  iron  c  fixed  on  the  frame.  This  rests  upon  the 
bottom  rock.  At  the  top  is  an  adjustable  pointed  round  iron  d  the  lower 
9  in.  of  which  is  threaded  so  that  by  means  of  the  nut  e  countersunk  as 
shown  in  the  frame,  when  a  wrench  is  applied  to  the  squared  portion 
above  the  thread  the  point  can  be  forced  up  against  the  roof  and  the 
frame  thus  held  securely  in  place.  A  3/8-in.  wire  rope  is  wound  two 
or  three  times  around  the  wheel  a  so  as  to  give  it  sufficient  grip  on  the 
wheel.  At  the  other  end  of  the  track  along  the  face  at  j  this  rope  passes 


82  MINING    WITHOUT    TIMBER 

through  an  ordinary  6-in.  iron  block  and  tackle  which  is  hooked  to  a 
chain  placed  around  the  prop.  One  end  of  the  rope  is  attached  to  the 
front  end  of  the  buggy  and  the  other  to  the  back  end  of  the  buggy.  By 
turning  the  handle  /  of  the  wheel,  the  buggy  can  be  moved  forward  and 
backward  along  the  face.  This  buggy  is  made  of  timber,  holds  about 
20  cu.  ft.,  and  one  side  is  20  in.  high  and  the  other  18  in.,  this  difference 
being  made  to  allow  room  for  loading  over  the  side.  The  wheels  are 
6  in.  in  diameter  placed  on  1  1/4-in.  axles.  In  the  bottom  there  is  an 
iron  plate  which  slides  in  and  out  sideways,  being  moved  by  a  handle. 
The  track  h  along  the  face  A  which  is  just  being  started  extends  out  over 
the  track  e  and  by  pulling  out  the  slide  in  the  bottom  of  the  buggy  the 
coal  is  dumped  from  the  buggy  into  the  mine  car  standing  on  the  track  e. 

The  coal  is  not  undercut,  and  in  general  does  not  need  to  be  drilled 
or  blasted,  the  weight  of  the  cover  being  generally  sufficient  to  loosen 
the  coal  with  an  occasional  shot  when  the  roof  pressure  is  not  sufficient. 
Owing  to  the  broken  conditions  of  the  measures  due  to  the  mining  out  of 
the  underlying  seams  the  coal  in  many  places  is  loose  and  simply  needs  to 
be  picked  out.  Six  men  work  along  each  face,  three  miners  and  three 
laborers  who  load  the  coal  into  the  buggies.  The  face  is  worked  in  several 
sections  as  shown  at  B,  each  section  being  taken  out  for  a  certain  distance, 
about  12  yd.  depending  upon  the  ease  with  which  the  coal  can  be  braken 
down;  but  no  section  of  one  face  is  allowed  to  get  far  ahead  of  any 
other  no  matter  how  easily  the  coal  can  be  mined.  By  the  time  section 
3  has  been  mined  out  the  coal  in  section  1  will  have  again  loosened  by 
the  weight  of  the  cover  and  can  then  be  taken  out  after  the  cogs  have 
been  built.  The  track  h  is  moved  near  the  face  after  each  section  is 
mined  and  a  row  of  cogs  is  kept  close  up  to  the  track.  Each  face  of 
coal  is  kept  about  40  ft.  in  advance  of  the  next  following  face.  The 
cogs  are  6  ft.  square  and  the  rows  are  8  ft.  apart  parallel  to  the  face  and 
12  ft.  apart  perpendicular  to  the  face.  These  cogs  are  built  from  the 
slate  overlying  the  coal,  and  as  it  comes  down  in  large  slabs  a  very  firm 
cog  is  formed.  The  space  between  the  large  rocks  is  filled  in  with  dirt 
and  a  perfectly  solid  cog  thus  formed. 

As  already  noted,  after  the  gateways  e  have  been  driven  in  full  height, 
a  distance  sufficient  to  allow  the  mine  car  to  be  placed  in  the  gateway  out 
of  the  way  of  the  haulage  on  the  main  gangway,  the  height  of  the  gate- 
way is  decreased  to  the  thickness  of  the  seam  and  overlying  slate,  that  is, 
about  36  in.  The  coal  is  moved  from  the  face  to  the  car  at  the  mouth  of 
the  gateway  by  means  of  a  buggy  similar  to  that  used  along  the  face  and 
already  described,  but  instead  of  the  winch  shown  in  Fig.  120,  it  is  moved 
by  means  of  a  double  winch  placed  at  the  point  h,  Fig.  1 19,  on  the  gateway 
where  the  height  is  decreased  to  the  height  of  the  seam.  This  winch, 
Fig.  121,  has  two  drums  a  and  a'  which  run  loosely  on  the  axle  b,  but  by 
means  of  the  clutches  c  and  c'  by  means  of  a  lever  not  shown,  either  drum 


ADVANCING    LONGWALL    SYSTEMS    FOR    SEAMS 


83 


may  be  made  to  turn  when -the  handle  d  is  moved.  If  the  load  is  too 
great  to  be  moved  by  turning  the  handle  d  the  winch  may  be  operated  on 
second  motion  by  means  of  a  pinion  e  attached  to  a  movable  axle  /. 
One  box  of  this  axle  at  g  is  loosely  bolted  to  the  framework  allowing  a 
little  play  of  the  axle,  while  in  box  h  is  an  elliptical  instead  of  a  circular 
hole  through  which  axle  /  passes  so  that  it  can  be  pushed  over  to  the 
dotted  position  /'  throwing  pinion  e  out  of  gear.  The  winch  is  set  on  a 
framework  of  timbers  one  end  of  which  rests  directly  on  the  bottom, 
while  the  other  end  is  let  into  a  groove  in  a  prop.  There  are  two  ropes 


FIG.  121. — Double  winch,  Bull's  Head  colliery. 

which  wind  upon  the  drums  a  and  a'.  One  of  these  is  attached  to  one 
end  of  the  buggy,  while  the  other  passes  to  the  upper  end  of  the  gateway, 
thence  through  a  small  6-in.  iron  block  and  tackle  k  fastened  to  a  prop 
by  a  chain  and  back  to  the  other  end  of  the  buggy,  or  by  means  of  guide 
pulleys  or  rollers  at  the  inby  end  of  the  gateway  the  rope  may  be  carried 
along  the  face  to  a  return  pulley  m  at  the  extreme  end  of  the  face  and 
the  gateway  buggy  taken  along  the  face  and  the  coal  brought  directly 
from  the  face  to  the  gangway. 

The  conditions  for  operating  the  longwall  system  of  mining  are  par- 
ticularly unfavorable,  for  the  bottom  is  badly  broken  and  a  stable  footing 
for  props  is  often  unobtainable.  At  the  face  of  one  of  the  gateways  at 
the  time  of  our  visit  the  bottom  had  dropped  away  entirely  from  beneath 
the  coal  leaving  the  coal  supported  only  by  contact  with  the  overlying 


84  MINING    WITHOUT    TIMBER 

slate.  The  top  is  also  badly  broken,  allowing  the  surface  water  to  enter 
the  mine  and  giving  a  roof  that  cannot  be  controlled.  This  roof  settles 
down  over  the  gangway  packs  about  2  ft.  so  that  while  the  gangway  is 
driven  about  6  ft.  high  it  is  only  about  4  ft.  high  after  the  workings  have 
settled.  Over  the  gateways  and  cogs  the  roof  settles  about  2  ft.  Thus 
far  an  output  of  1800  tons  per  foot-acre  has  been  obtained,  a  much 
better  yield  than  is  usually  obtained  in  anthracite  mining. 

EXAMPLE  52. — VINTON    BITUMINOUS    COLLIERY,    VINTONDALE,    PENN. 

Thin  Sloping  Seam,  800  ft.  Deep;  Panel  System;  Loading  into 
Pan  Conveyors. — Transporting  coal  from  the  working  face  to  main 
haulage  roads  by  means  of  mechanical  conveyors  is  a  comparatively 
recent  departure  from  ordinary  mining  methods.  This  system,  which 
was  first  introduced  in  England,  was  early  recognized  by  leading  opera- 
tors there  as  possessing  superior  advantages  over  the  usual  manner  of 
working,  especially  in  thin  coal  seams  where  the  roof  has  to  be  brushed 
to  allow  all  but  the  tiniest  cars  to  reach  the  long  wall  face. 

The  coal  worked  at  Vintondale  is  bound  tight  to  roof  and  floor  and 
is  the  "  B "  or  Lower  Kittanning  seam,  42  in.  in  thickness,  which  lies 
on  a  pitch  of  8  per  cent.,  with  an  average  of  200  feet  of  cover.  The  coal 
is  of  a  soft  and  friable  nature,  free  from  slate  bands  and  bony  coal,  but 
interspersed  with  sulphur  pyrites,  which,  at  times,  cause  considerable 
annoyance  in  cutting  and  drilling.  The  bottom  is  a  mixture  of  coal  and 
fireclay,  while  the  roof  is  composed  of  from  8  to  12  ft.  of  black  slate, 
overlaid  with  sandstone.  The  slips  in  the  slate  are  well  marked,  and 
lie  at  an  angle  of  25  deg.  with  the  line  of  greatest  dip ;  the  longwall  face 
is  kept  normal  to  these  slips.  The  present  panel  modification  of  longwall 
mining  was  first  started  in  No.  3  mine  in  1900.  At  the  outset  cars  were 
run  around  the  working  face  and  loaded.  This  method  brought  only 
fair  results,  owing  to  the  necessity  of  using  small  cars,  steep  grades,  and 
difficulty  in  keeping  roadways  open. 

Arrangements  were  then  made  for  the  placing  of  a  conveyor  along  the 
face,  allowing  the  cars  to  be  run  under  the  head-end  to  be  loaded.  The 
first  conveyor,  which  was  made  entirely  of  wood,  was  a  cumbersome 
affair,  and  much  time  was  consumed  in  moving  it  laterally  along  the 
face  after  the  cut  had  been  loaded  out;  but,  after  a  year's  trial,  the 
results  obtained  were  so  gratifying  that  metal  conveyors  were  designed 
and  ordered,  and  preparations  were  made  to  employ  this  system  on  a 
much  larger  scale  (Fig.  122). 

The  metal  conveyor  consists  of  a  trough  or  pan,  made  of  sheet  steel 
1/8  in.  thick,  12  in.  wide  at  the  bottom,  18  in.  wide  at  the  top,  and  6  in. 
high,  set  on  strap-iron  standards  as  shown  in  detail  in  Fig.  123.  A  con- 
veyor is  made  up  in  sections  of  6-,  12-,  15-,  and  18-ft.  lengths,  connected 


ADVANCING    LONGWALL    SYSTEMS    FOR    SEAMS 


85 


together  by  means  of  1/2-in.  flatheaded  bolts,  countersunk.  The  front 
is  inclined  for  a  distance  of  45  ft.  to  allow  clearance  for  mine  cars  to  pass 
under  (see  Fig.  114).  The  rear  end  is  inclined  for  15  ft.  to  compensate 
for  the  size  of  sprocket  wheel.  A  return  runway  for  the  chain  is  afforded 
below  the  pans  by  angle  irons. 

A  cast-iron  driving  sprocket,  18  in.  in  diameter  and  13-in.  face,  is 


Section  C-D 
Pia.  122. — Plan  and  section  of  Vinton  conveyor  system  showing  head  of  main  conveyor. 

attached  to  the  front  end.  On  the  shaft  of  this  sprocket,  which  is 
extended  12  in.  beyond  one  of  the  bearings,  is  keyed  a  12-tooth,  16-in. 
diameter  sprocket,  which  connects  with  the.  driving  mechanism.  The 
rear-end  section  (c)  consists  of  a  framework  made  up  of  two  I  beams,  6  ft. 
long  and  strongly  braced,  on  which  rest  the  take-up  boxes  for  keeping 
the  chain  in  adjustment,  and  the  rear  sprocket  wheel  over  which  the 


Pia.  123. — Cross  section  of  conveyor,  Vinton  colliery. 

chain  returns.  There  are  two  conveyor  chains,  held  apart,  the  width  of 
the  trough,  by  crossbolts  which  act  as  scrapers  to  replace  the  usual 
plates.  The  chains  are  of  steel  and  are  designed  for  quick  repairing. 

The  triple  conveyor  system  (Fig.  124),  was  finally  designed  and  in- 
stalled as  an  improvement  over  the  single  type.  In  laying  out  a  mine 
for  this  system,  the  main  entry  and  airway  are  driven  up  or  down  the 


86  MINING    WITHOUT   TIMBER 

pitch,  and  cross-headings  are  driven  off  them  at  intervals  of  400  ft.  at 
such  an  angle  as  will  give  a  2-per-cent.  grade;  75-ft.  barrier  pillars  are  left 
on  each  side  of  the  main  entries.  The  cross-heading  is  driven  20  ft.  wide 
and  gobbed  on  the  lower  side.  The  air-course,  which  afterward  is  used 
as  the  panel  or  block  face,  is  driven  20  ft.  wide,  but  no  bottom  is  lifted; 
a  40-ft.  pillar  is  maintained.  Block  headings  are  run  perpendicular 
to  cross-headings  at  518-ft.  centers;  they  are  driven  18  ft.  wide,  with 
bottom  lifted  in  the  center  5  ft.  wide,  and  deep  enough  for  a  5-ft. 
clearance. 

When  the  block  is  ready  for  operation,  a  conveyor  350  ft.  long  is 
placed  in  the  block  heading,  and  along  the  face  of  the  air-course  on  each 


Fia.  124. — Plan  triple  conveyor  system,  Vinton  colliery. 

side  is  placed  a  conveyor  250  ft.  long,  with  delivery  ends  directly  over 
the  main  conveyor,  one  being  5  ft.  in  advance  of  the  other.  Each  con- 
veyor is  driven  by  a  20-horsepower,  250-volt,  series-wound  motor,  en- 
cased in  a  sheet-iron  frame  mounted  on  steel  shoes,  so  as  to  be  easily 
moved. 

Airways  are  maintained  on  the  blocks  by  driving  two  places  slightly 
in  advance  of  the  block  face,  6  and  4  ft.  wide,  respectively,  with  a  10-ft. 
pillar  between.  The  first  place  acts  as  a  stable  for  the  machine,  and  is 
driven  by  the  machine.  The  airway  is  pick-mined,  and  one  man  manages 
to  keep  these  places  going  on  the  rear  end  of  both  blocks.  By  this  ar- 
rangement no  cribbing  is  necessary. 

The  blocks  are  worked  to  within  25  ft.  of  the  cross-heading,  when 


ADVANCING    LONGWALL   SYSTEMS    FOR   SEAMS  87 

the  conveyors  are  removed  to  another  block.  The  remaining  pillar  is 
brought  back  along  with  the  heading  stumps. 

The  power  is  carried  to  the  top  of  the  block  heading  by  a  00  wire. 
Here  are  attached  two  insulated  twin  cables,  one  to  furnish  power  to  the 
machines,  the  other  for  the  drives  and  hoist. 

The  cables  are  carried  down  the  block  heading,  one  on  each  side  of 
the  main  conveyor,  being  attached  to  it  by  means  of  malleable-iron 
brackets.  At  the  junction  of  the  conveyors  connections  are  made  with 
the  drives,  also  with  a  cable  that  is  attached  to  each  of  the  face  conveyors. 

Stations  are  established  50  ft.  apart  on  the  face  conveyor  cables,  to 
which  connections  are  made  with  the  short  cable  attached  to  the  longwall 
machines  and  electric  drills.  Switches  are  placed  at  the  head-end  of  the 
main  conveyor,  by  which  the  power  is  controlled. 

The  method  of  handling  the  cars  to  the  conveyor  is  simple.  A  side 
track  is  laid  300  ft.  long,  of  which  the  block  heading  is  the  center.  Con- 
nection is  made  with  the  main  track  at  the  lower  end,  and  a  cross-over 
switch  is  placed  directly  under  the  conveyor.  At  the  upper  end  of  the 
siding  is  placed  an  electric  hoist.  A  trip  of  14  cars  is  shoved  into  the 
empty  track,  and  the  rope  is  attached  and  the  trip  pulled  up  to  the  con- 
veyor. Signal  wires  are  hung  between  the  conveyer  and  the  hoist,  and 
and  as  each  car  is  loaded  the  trip  is  pulled  forward.  When  loaded, 
trip  is  dropped  on  the  loaded  siding,  the  rope  disengaged  and  attached 
to  the  empties. 

The  crew  operating  a  double  block  consists  of  17  men,  i.  e.,  block 
boss,  machine  runner  and  helper,  driller,  shooter,  two  conveyor  men, 
hoist  boy,  five  loaders,  and  four  timbermen.  Two  longwall  machines  of 
Jeffrey  or  Sullivan  make  are  used,  one  for  each  side,  although  one  ma- 
chine can  keep  up  the  work  in  case  of  emergency.  The  machine  men 
finish  cutting  one  block,  in  five  hours,  and  then  put  the  machine  in 
position  to  start  back  on  the  cut  and  move  over  to  the  next  block  and 
begin  cutting.  They  are  followed  by  the  shooter  and  loaders. 

When  a  block  is  cleaned  up,  the  timbermen  move  up  the  conveyor, 

This  consists  of  setting  a  line  of  props,  called  the  line  row,  about  8  ft. 
apart,  and  a  distance  from  the  conveyor  equal  to  the  depth  of  the  under- 
cut. As  these  are  placed  the  old  line  row,  which  is  now  against  the  con- 
veyor, is  withdrawn.  The  pulling  jacks  for  moving  the  conveyor  are 
distributed  along  the  block  40  ft.  apart  and  placed  in  position. 

The  shot  firer  keeps  closely  after  the  machine,  and  is  through  shooting 
shortly  after  the  undercut  is  finished.  The  driller  then  starts  from  the 
far  end  of  the  block  to  drill  holes  in  the  new  face.  It  usually  takes  him 
about  two  hours  to  drill  the  entire  width  of  the  block. 

Each  loader  is  supplied  with  a  pick  and  shovel  and  a  piece  of  sheet 
iron  9  in.  wide  and  6  ft.  long,  which  he  attaches  to  the  conveyor  to  act 
as  a  sideboard.  As  each  loader  cleans  up  his  place  he  moves  forward  to 


88  MINING    WITHOUT   TIMBER 

the  head  of  the  line.  This  continues  until  the  coal  is  loaded  out,  which 
usually  requires  about  six  and  a  half  hours. 

When  cleaned  up,  the  drive  is  reversed  and  the  timber  which  has 
arrived  on  the  last  trip  is  run  through  on  the  conveyor  to  such  points  on 
the  block  where  it  is  required.  When  this  is  accomplished  the  power 
is  shut  off,  and  the  conveyor  is  moved  up  to  the  line  row.  This  lateral 
move  of  the  conveyor  requires  very  little  time,  very  seldom  exceeding  five 
minutes.  A  break  row,  consisting  of  two  rows  of  props  set  2  ft.  apart,  is 
now  placed  along  the  lower  side  of  the  conveyor.  These  props  are  set 
on  a  cap  piece,  placed  on  a  small  pile  of  slack,  and  wedged  at  the  top. 
Two  break  rows  are  all  that  is  necessary  to  protect  the  block.  In  the 
meantime,  a  portion  of  the  crew  are  engaged  in  pulling  out  the  extra 
break  row.  This  is  the  most  hazardous  work  on  the  block,  and  is  given 
personal  attention  by  the  block  boss.  'Axes  are  used  in  this  operation, 
and  about  75  per  cent,  of  the  props  recovered  are  practically  uninjured. 

While  part  of  the  crew  are  employed  timbering,  the  rest  make  the 
necessary  connections,  and  go  along  the  conveyor  with  a  pump  jack  and 
level  it  up.  They  also  build  a  crib  at  the  head  end,  which  is  placed  to 
prevent  the  roof  from  breaking  over  into  the  block  heading.  All  the  dead 
work  is  taken  care  of  by  the  four  timbermen,  thus  not  hindering  the 
steady  flow  of  coal,  which  averages  150  tons  daily,  from  a  5-ft.  undercut. 

For  the  purpose  of  keeping  the  machinery  in  as  good  shape  as  pos- 
sible, a  skilled  mechanic  is  attached  to  each  mine.  He  assumes  charge 
in  case  of  an  accident  and  makes  necessary  repairs,  although  most  of  the 
breakdowns  are  easily  taken  care  of  by  the  block  boss  and  machine  man. 

In  the  starting  of  a  block  is  where  the  best  results  are  obtained,  as 
the  roof  requires  little  attention  until  about  100  ft.  have  been  extracted. 
It  then  begins  to  weigh  heavy  on  the  posts,  and  it  is  found  necessary  to 
carry  three  or  four  double  break-rows  with  cogs  in  anticipation  of  what 
is  called  the  "big  break."  This  usually  occurs  when  the  block  is  ad- 
vanced from  100  to  150  ft.,  although  in  several  instances  a  500-f  t.  face 
has  been  carried  up  200  ft.  before  the  overhanging  strata  broke.  After 
the  sand  rock  is  down,  only  two  break  rows  are  carried,  and  the  roof 
keeps  breaking  behind  the  last  row  as  the  face  is  extended. 

The  men  are  paid  day  wages  and,  as  they  become  accustomed  to  the 
work  and  machinery,  are  advanced  accordingly.  The  "block"  boss, 
as  an  incentive  to  secure  the  best  results,  is  paid  a  small  bonus  per  ton 
besides  his  regular  day  rate.  The  cost  averages  for  the  last  two  years 
show  that  block  coal  is  loaded  on  the  mine  cars  35  per  cent,  cheaper  than 
the  district  mining  rate  for  pick  work  with  loading  into  cars. 

The  above  conditions  prevailed  in  Nov.,  1907,  but  on  the  author's 
visit  in  Sept.,  1910,  he  found  the  longwall  system  superseded  by  the 
former  room  and  pillar  system  for  the  following  assigned  reasons.  1. 
The  frequent  breakage  of  the  conveyors  caused  a  very  irregular  output. 


ADVANCING    LONGWALL    SYSTEMS    FOR   SEAMS  89 

2.  The  miners  preferred  the  contract  payments  of  room  and  pillar  to  the 
time  wages  of  the  longwall  system.  3.  The  timber  consumption  was 
excessive,  because  many  of  the  props  could  not  be  recovered. 

None  of  these  disadvantages  are  irremediable,  and  the  cost  of  timber 
can  always  be  obviated  wherever  enough  slate  can  be  cheaply  got  from 
the  floor,  parting,  or  roof  to  build  pack  walls  to  replace  part  of  the  props 
and  cogs.  The  conveyor-longwall  system  has  proved  profitable  for  thin 
seams  in  Europe,  and  the  Vintondale  method  should  prove  commercially' 
successful  in  other  American  fields  where  the  natural  conditions  are 
suitable. 

EXAMPLE  53. — DRUMMOND    BITUMINOUS  COLLIERY,  WESTVILLE,   N.  S. 

Thick  Sloping  Seam  at  2000-/Y.  Depth.  Loading  into  Cars  handled 
on  "Jigs." — When  coal  workings  extend  beyond  a  vertical  depth 
of  1500  ft.,  it  generally  becomes  unprofitable,  if  not  impossible,  to 
work  by  one  of  the  "pillar"  methods,  for  the  enormous  weight  of  the 
overlying  strata  will  not  only  break  and  crush  the  timber,  but  will  also 
either  crush  the  pillars  or  force  them  into  the  strata  immediately  above 
or  below  the  seam,  resulting  in  a  " creep"  and  the  closing  up  of  roads. 

The  size  of  the  pillars  must  increase  with  the  depth,  until  at  about 
the  depth  noted  above,  the  pillars  become  so  large,  and  the  amount  of 
coal  that  can  be  safely  worked  so  small,  especially  if  it  is  of  a  friable 
nature,  that  the  operations  become  unprofitable,  and  another  method 
must  be  adopted  or  the  mine  closed  up. 

Such  was  the  situation  in  1896  at  this  mine  in  working  by  the  room 
and  pillar  method  a  17-ft.  seam  of  friable,  gaseous  coal,  with  a  very  weak 
roof  of  black  carboniferous  shale,  the  seam  dipping  from  18  to  27  deg. 

The  mine  had  been  developed  by  two  parallel  slopes  on  the  dip,  and 
from  these  double-entry  "lifts"  were  turned  off  every  400  ft.  These 
entries  or  levels  were  9  ft.  wide  by  7  ft.  high,  and  as  depth  was  gained  it 
was  found  difficult  to  support  their  roofs.  At  a  depth  of  1200  ft.,  the 
first  "chocks"  or  cogs  had  to  be  built  'on  each  side  of  the  level.  In  the 
next  lift,  400  ft.  below,  it  became  necessary  to  change  the  working  system 
if  the  coal  were  to  be  mined  at  a  profit.  The  change  was  made  without 
great  expense,  any  interruption  of  the  regular  output  or  any  considerable 
variation  in  the  ventilation,  etc. 

The  advancing  panel  system  of  longwall  was  adopted  and  it  has  proved 
quite  successful  considering  the  depth  reached,  which  is  7,870  ft',  on  the 
slopes  or  over  2000  ft.  vertically.  The  new  system  has  been  partic- 
ularly free  from  fatal  accidents  at  the  face,  those  occurring  happening 
in  the  roadways,  etc.  The  slopes  are  sunk  as  formerly,  diverging  slightly 
to  increase  the  pillar  of  solid  coal  between  them.  They  are  supported 
on  either  side  by  pillars  also  increasing  in  width  with  depth  so  that  they 
are  now  about  350  ft.  wide. 


90 


MINING    WITHOUT   TIMBER 


Either  one  or  both  of  these  slopes  is  used  as  the  intake  airway,  Fig. 
125,  while  return  airways  are  maintained,  one  on  each  side  along  the  slope 
pillars.  Two  levels  are  driven  as  formerly  which  form  a  lift  with  about 
400  ft.  of  solid  coal  between  pairs  of  levels.  The  upper  level  of  each  pair 
is  used  as  a  haulage  road,  and  the  lower  level  forms  the  intake  airway 
for  each  lift.  This  intake  carries  fresh  air  from  the  slope,  where  it  is 
split,  to  the  inner  workings  first;  from  there,  returning  and  ascending,  it 
passes  through  each  of  the  working  places  to  the  lift  above,  and  thence  to 
the  return  airway;  it  is  also  used  for  drainage,  and  generally  there  is  a 
dam  built  on  it  near  the  slope  which  catches  all  the  water  from  the  lift. 


Fio.  125. — Plan  of  layout,  Drummond  colliery. 

The  levels  are  driven  as  nearly  parallel  as  possible,  rising  about  1  ft. 
in  130  ft.  with  from  15  to  20  ft.  of  solid  coal  between  the  chocks.  This 
pillar  is  often  removed  and  the  space  filled  in  with  stone  from  the  roof, 
the  result  of  "brushing"  which  must  be  done  very  shortly  after  the  levels 
are  driven.  These  levels  are  driven  8  ft.  wide  and  8  ft.  high,  they  are 
first  made  about  18  ft.  wide  and  7  ft.  high.  This  leaves  a  "bench"  on 
the  bottom  which  is  only  cut  in  the  case  of  roadways.  On  this  bench 
chocks  are  built  quite  close  together  on  each  side,  and  about  8  ft.  apart 
across  the  road,  Figs.  126  and  127,  with  sided  timber  over  them  across 
the  road  about  3  ft.  apart,  and  slabs  over  the  timber  to  support  the  roof. 
The  chocks  are  built  of  blocks  of  wood  over  5  in.  in  thickness  and  5  ft.  in 
length,  making  them  5  ft.  square.  After  these  are  built  (similar  to  logs 
in  a  wharf)  the  bench  is  cut  along  the  chocks  and  the  bottom  lifted  to 
give  the  8-ft.  height. 

Off  these  levels,  "jigs"  are  driven  up  on  the  full  pitch  of  the  coal, 
Figs.  127  and  128,  not  more  than  400  ft.  apart;  they  are  chocked  as  well 
as  all  other  roadways.  An  airway  5  ft.  wide  is  carried  up  on  the  side  of 
this  jig  farthest  from  the  slope,  and  the  chocks  on  this  side  must  be  made 


ADVANCING    LONGWALL    SYSTEM    FOR   SEAMS 


91 


air-tight.  This  is  done  by  filling  them  with  stone  and  fine  coal,  etc. 
Owing  to  the  very  heavy  pressure  required  in  maintaining  ventilation 
at  this  depth,  canvas  doors  can  only  be  used  as  a  temporary  arrangement. 
A  wooden  door  is  placed  in  an  air-tight  frame  across  the  level  to  direct 
the  air  up  this  airway  between  the  coal  and  the  air-tight  chocks,  passing 


FIG.  126. — Cross  section  of  road,  Drummond  colliery. 

around  the  face  and  returning  down  the  jig  which  is  8x8  ft.,  Fig.  127. 
This  practice  has  been  proved  many  times  to  be  the  only  practical  way, 
as  the  air  will  not  pass  up  the  large  and  down  the  smaller  airway  in 
sufficient  quantity  to  keep  the  face  clear  of  gas.  This  method  is  con- 
tinued until  the  jig  is  driven  through  to  the  lower  level  of  the  lift  above, 

A   Air  Tight  Chocks 
S    Stopping 


FIG.  127. — Plan  of  gateway  and  face,  Drummond  colliery. 

when  the  door  is  removed  and  the  air  passes  up  the  jig  and  out  to  the 
airway,  and  the  airway  along  the  chocks  is  allowed  to  cave. 

Working  the  Rooms. — Beginning  at  the  lower  entry  of  the  lift  above 
on  one  of  these  jigs,  rooms  are  broken  off  with  about  41  ft.  between  the 
centers.  The  "gateway"  or  road  in  the  rooms  is  much  the  same  as  the 
levels  already  described.  They  are  timbered  in  the  same  way,  Fig.  128, 


92 


MINING    WITHOUT    TIMBER 


except  that  the  chocks  are  built  about  2  ft.  apart  on  each  side,  and  only 
about  6  ft.  apart  across  the  road.  When  this  gateway  is  driven  in  about 
25  ft.,  work  is  started  on  the  breast.  From  it  the  coal  is  all  taken  out 
to  a  thickness  of  7  ft.  and  up  to  the  room  or  level  above.  The  breast  is 
then  timbered  with  upright  timber  props  with  cap-pieces  between  them 
and  the  roof.  Sometimes  sided  timbers  are  placed  with  one  end  on  the 
high-side  chock  of  the  roadway  and  props  under  the  middle  and  upper 
end.  The  gateway  is  kept  15  ft.  to  20  ft.  ahead  of  the  breast,  the  roof 
of  which  is  allowed  to  fall  in  as  the  face  advances;  generally  when  about 
40  ft.  from  the  jigs  the  roof  falls,  often  causing  a  great  smashing  of 
timber  on  the  road  below,  the  bottom  rising  up  as  well.  The  face  of  the 


FIQ.  128.— Plan  of  jig  road,  Drummond  colliery. 

gateway  is  kept  a  short  distance  ahead,  for  if  the  face  of  road  were  in 
line  with  the  face  of  the  breast,  it  would  be  very  apt  to  fall  solid  across  the 
face  of  the  road  as  well,  and  take  a  week  or  more  to  get  into  working 
shape,  again.  Through  carelessness  of  the  miners  this  sometimes  happens. 
No  explosives  are  used,  for  if  these  places  are  properly  timbered  and  the 
weight  thrown  on  the  face  the  coal  is  easily  worked  with  hand  picks,  but 
wedges  are  required  in  lifting  the  bench  in  the  roadways. 

^  Quite  often  the  roof  falls  in  solid  to  the  face,  then  it  is  necessary  to 
drive  a  heading  up  in  the  solid  coal  at  the  face  and  start  the  breast  over 
again.  This  perhaps  is  the  greatest  difficulty  met  with  in  the  whole 
operation,  for  when  a  fall  like  this  takes  place  the  ventilation  is  cut  off 
and  generally  some  gas  accumulates  and  when  the  heading  is  started  up, 


ADVANCING    LONGWALL    SYSTEMS    FOR   SEAMS  93 

the  gas,  also  rising,  follows  the  miner  and  causes  trouble  before  it  can  be 
driven  20  ft.  or  35  ft.  to  the  place  above. 

Generally  three  of  these  rooms  are  worked  on  each  side  of  the  jig 
simultaneously,  the  upper  ones  leading  and  the  others  following  in  step- 
like  order  from  20  ft.  to  40  ft.  behind  the  preceding  one.  In  this  way 
the  upper  7  ft.  of  the  17-ft.  thickness  of  coal  in  the  seam  is  taken  out  in 
one  operation  and  in  future  years  the  balance  may  be  mined  similarly. 

Three  miners  and  a  laborer  work  in  each  room,  as  the  success  of  this 
method  requires  that  the  breasts  be  kept  steadily,  if  slowly,  advancing. 
With  depth  it  is  necessary  to  shorten  the  gateways  in  order  that  the  coal 
may  be  all  taken  out  before  they  become  entirely  closed  up,  for  on  every 
side  may  be  seen  examples  of  both  squeeze  and  creep.  The  roof  pres- 
sure is  so  great  that  thin  clay  partings  in  the  coal  squeeze  out  like  clay 
from  a  brick-machine,  while  the  lateral  pressure  on  the  coal  walls  re- 
duces the  space  of  the  openings  30  per  cent,  in  a  few  months.  The 
combined  pressures  so  break  ordinary  booming  in  about  a  month  that  it 
becomes  necessary  to  brush  and  retimber  again.  Here  places  may  be 
seen  so  closely  timbered  that  neither  rock  nor  coal  are  to  be  seen  for 
long  distances  except  at  the  working  faces.  Studies  are  constantly 
made  of  the  faults,  cleats  and  clay  partings  encountered,  (so  as  to  make 
the  pressure  mine  the  coal  with  the  least  labor),  of  the  proper  setting 
of  timbers,  of  the  breaking  away  of  the  ribs  and  how  to  avoid  it;  and  of 
how  to  distinguish  the  actual  sounds  of  danger  as  there  is  always  some 
cracking  of  timbers  heard  in  the  working  places.  So  expert  do  those 
extracting  the  coal  from  the  breast  become  that  they  work  on  to  the 
last  minute  before  a  fall  takes  place  amid  appalling  conditions. 

Haulage. — There  is  no  mechanical  haulage  on  the  levels,  the  work 
being  done  by  horses.  At  the  bottom  of  the  jigs  and  at  the  mouths  of 
the  rooms — which  are  opposite  each  other,  three  on  each  side  of  the  jig — 
large  metal  plates  are  laid  on  timbers  placed  horizontally  and  made 
solid  in  that  position.  When  laid  they  form  a  smooth  surface  from  6  to 
8  ft.  square.  On  these  the  cars  can  easily  be  turned  in  any  direction. 

The  road  on  the  jig  is  either  a  double  track  or  three  rails,  with  a 
passing  turnout  halfway  (see  Fig.  128).  Over  the  two  lower  plates  on 
the  jig,  lifting  rails  about  8  ft.  long  are  fitted  into  clips.  When  running 
coal  from  the  lower  rooms,  these  rails  are  removed  and  a  tail-rope  the 
necessary  length  is  attached  with  a  safety  hook.  A  drum,  controlled  by  a 
boy,  is  placed  at  the  top  and  the  weight  of  the  full  car  running  down  takes 
the  empty  car  up.  The  cars  run  on  their  own  wheels  from  the  surface 
to  the  face,  their  gauge  is  2  1/3  ft.,  and  their  capacity  is  1600  Ib.  of  coal. 

Output  and  Timber  Used. — No  timber  is  drawn,  the  great  difficulty 
being  to  get  enough  timber  in  to  keep  sufficient  room  for  the  proper 
ventilation  of  the  mine.  The  mine  produces  about  1200  tons  of  coal 
a  day  and  consumes  two  thousand  5-ft.  sticks,  but  of  a  small  diameter. 


CHAPTER  XX 
PILLAR  SYSTEMS  FOR  SEAMS 

EXAMPLE  54. — ADVANCING-SYSTEM  LAYOUTS  FOR    ROOM  AND  PILLAR, 
PILLAR  AND  STALL,  AND  PANEL  METHODS 

Room  and  Pillar  System. — The  pillar  system  is  also  known  as 
"room  and  pillar,"  "pillar  and  chamber,"  "bord  and  pillar,"  etc.  It 
is  applicable  to  all  classes  and  conditions  of  mining  where  the  roof  pressure 
is  not  such  as  to  destroy  pillars  of  reasonable  sizes,  subject,  however,  to 
such  modifications  as  serve  to  adapt  it  to  the  varying  conditions  of  weak 


Fia.  129. — Layout  for  Room  and  Pillars  system,  flat  seam. 

or  strong  roof  or  floor;  tough,  friable,  or  gaseous  coal;  predominance  of 
face  or  end  cleats;  inclination  of  the  seam,  etc.  The  features  of  the  sys- 
tem are  openings  driven  square  from  or  at  an  angle  to  the  haulway. 
Such  opening  may  be  driven  wide  or  narrow,  and  may  be  a  roadway, 
incline,  or  chute,  as  best  adapted  to  the  existing  conditions. 

Room  and  Pillar  System  (proper)  .—This  layout  as  applied  to 
flat  seams,  or  where  the  inclination  does  not  exceed  3  deg.,  is  illustrated 
in  Fig.  129.  The  shaft  bottoms,  including  the  stable,  are  here  shown 

94 


PILLAR   SYSTEMS    FOR    SEAMS 


95 


crossing  the  shaft  pillar  at  an  angle  conforming  to  the  surface  tracks, 
thereby  giving  a  straight  dump  and  tipple  in  line  with  the  shaft.  The 
stables  are  located  close  to  the  shaft  bottom,  where  the  mules  can  be 
rescued  in  case  of  accident,  and  where  the  daily  feed  and  refuse  can  be 
conveniently  handled.  Free  access  is  had  to  the  stables  from  the  main 
haulage  roads  without  passing  through  a  door;  while  immediate  access  is 
had  close  to  the  shaft  through  a  curtain  or  canvas.  Good  ventilation  is 
secured  by  a  small  separate  split  of  fresh  air,  while  the  return  air  from 
the  stables  at  once  enters  the  return  from  the  mine  and  passes  up  the 
shaft  without  contaminating  the  mine  air.  Another  feature  of  the 
arrangement  shown  in  Fig.  129,  is  the  small  number  of  doors.  The  coal 
coming  from  any  room  upon  the  main  road  of  a  pair  of  entries  has  no 


Fia.  130. — Layout  for  Room  and  Pillar  system,  sloping  seam. 

doors  to  pass  through;  while  that  coming  from  the  back  entry  of  each 
pair  has  but  one  door  to  pass  through  on  its  way  to  the  shaft.  This  is  a 
great  saving  of  expense  and  trouble  and  may  often  avert  possible  disaster 
arising  from  the  derangement  of  the  ventilation  by  doors  being  left  open. 
The  chambers  or  rooms  are  here  turned  square  with  the  entry  narrow 
for  a  distance  of  4  or  5  yards  and  then  widened  out  inbye,  the  road  in 
each  room  following  the  straight  rib.  The  waste  from  the  seam  is  stored 
in  the  room.  The  rooms  are  spaced,  under  normal  conditions  of  roof 
and  floor,  from  40  to  45  feet  apart,  center  to  center.  The  breast  is  usually 
8  yards  wide  and  is  driven  up  from  60  to  100  yards.  When  the  breast  is 
abandoned  the  miner  starts  to  draw  back  his  pillar  unless  for  special 
reasons  this  is  delayed  for  a  while. 


96 


MINING    WITHOUT    TIMBER 


In  Fig.  130  is  shown  the  application  of  the  room  and  pillar  system  to 
seams  pitching  from  3  to  5  deg.  It  differs  from  the  method  shown  in 
Fig.  129  by  turning  rooms  to  the  rise  only.  When  the  pitch  of  the  seam 
is  from  5  to  10  deg.,  the  car  may  still  be  taken  to  the  face  and  loaded  by 
driving  the  rooms  across  the  pitch,  or  at  an  angle  with  the  level  or  gang- 
way. This  reduces  the  grade  of  the  track  in  the  rooms.  When  the 
inclination  of  the  seam  is  still  greater,  buggies  are  sometimes  used,  the 
track  being  built  upon  the  refuse  of  the  seam  and  raised  at  its  lower  end 
where  a  tip  is  arranged  by  which  the  coal  is  dumped  from  the  buggy  into 
the  mine  car  ready  to  receive  it.  Where  the  coal  is  soft,  this  method 
cannot  be  used.  It  is  employed  on  pitches  not  exceeding  15  or  18  deg. 
in  thick  seams. 

Seams  pitching  more  than  15  deg.  are  usually  worked  by  chutes 
or  self-acting  inclines.  When  the  pitch  is  less  than  30  deg.  sheet  iron  is 


Fia.  131. — Room  and  Pillar  system  for  steep 


usually  laid  in  the  chute  as  a  floor,  to  enable  the  coal  to  slide  more  easily; 
but  on  inclinations  of  less  than  20  deg.  it  is  usually  necessary  to  push  the 
coal  down  the  chute  by  hand  or  by  mechanical  means,  as  it  does  not  slide 
readily.  On  pitches  steeper  than  30  deg.  sheet  iron  is  not  necessary,  as 
the  coal  will  slide  without.  Fig.  131  shows,  in  plan  A  and  section  B, 
the  arrangement  of  the  breasts,  chutes,  and  manways,  and  the  position 
of  the  gangway  and  air-course  at  the  roof  of  the  seam,  in  thick,  steep- 
pitching  anthracite  beds.  This  position  of  the  gangway  and  air-course 
secures  a  better  inclination  of  the  loading  chute  and  manways,  and  pre- 
sents less  danger  from  squeeze.  In  the  figure,  g  is  the  gangway  and  m 
the  manway  leading  to  the  dividing  at  the  floor  of  the  seam  into  two 
branches  s,  s,  which  lead  to  either  breast.  At  this  point,  or  slightly 
below  it,  a  small  cross-cut  d  is  driven  up  to  the  airway  c.  This  is  brat- 
ticed  off  and  used  only  in  case  of  need,  as  the  air  is  regularly  conducted 


PILLAR   SYSTEMS    FOR   SEAMS 


up  one  breast  manway  and  down  the  other  side  to  the  highest  cross-cut 
and  thence  to  the  next  breast.  Brattices  with  small  doors  are  also 
placed  in  the  manways  to  keep  the  air  from  taking  a  short  circuit  through 
the  manways.  Small  manways  are  bratticed  off  the  side  of  each  loading 
chute  for  the  use  of  the  loaders. 

Self-acting  inclines  are  used,  sometimes,  upon  steep  pitches  in  pref- 
erence to  chutes.  In  this  case,  butt  headings  are  usually  driven  to  the 
full  rise  and  rooms  set  off  on  the  strike  from  these  rise  headings,  buggies 
being  used  in  the  rooms  to  convey  the  coal  from  the  face  to  the  incline. 
It  is  hardly  necessary  to  state  that  dip  inclines  are  rarely  ever  introduced 
as  a  permanent  feature,  it  being  better  to  sink  the  main  slope  far  enough 
to  permit  another  level  from  which  the  coal  can  be  worked  to  the  rise. 


Fia.  132. — Single  Stall  system. 


FIG.  133.— Double  Stall  system. 


Stall  and  Pillar  System. — This  is  similar  to  the  system  just  described, 
except  in  the  relative  size  of  pillars  and  breasts.  It  is  adapted  to  weak 
roof  and  floor,  or  strong  roof  and  soft  bottom,  to  a  fragile  coal,  or  to 
other  similar  conditions  requiring  ample  support.  The  stall  system  is 
particularly  useful  in  deep  seams  where  the  roof  pressure  is  great.  The 
stalls  are  usually  opened  narrow  and  widened  inside  to  furnish  a  breast 
which  varies,  according  to  conditions  of  roof,  floor,  coal,  depth,  etc., 
from  4  to  6  yd.  wide  in  the  "single-stall"  method.  The  pillars  between 
the  stalls  are  usually  about  the  width  of  the  breasts. 

Fig.  132  shows  the  method  by  single  stall  and  Fig.  133  that  by  double 
stall.  The  former  is  more  applicable  to  flat  seams  or  seams  of  small 
inclination;  while  the  latter  is  used  on  steep  pitches.  The  single-stall 
method  affords  but  one  road  to  a  breast;  and,  hence,  does  not  permit  of 


98  MINING    WITHOUT   TIMBER 

the  concentration  of  men  possible  in  double  stalls  where  there  are  two 
roads  to  each  breast.  In  the  double  stalls  the  breasts  are  wider,  ranging 
from  12  to  15  yd.;  while  the  pillars  sometimes  reach  a  width  of  30  yd. 

Panel  System. — It  is  advisable  to  mine  in  panels:  1,  When  the  seam 
contains  much  gas,  making  it  essential  that  the  ventilation  of  the  entire 
mine  be  under  absolute  control;  2,  when  the  coal  is  readily  affected  by 
the  air,  and  disintegrates  with  long  standing;  3,  When  the  roof  pressure 
or  the  conditions  of  the  roof  are  such  as  to  require  extreme  caution  to 
prevent  squeeze  or  creep.  The  panels  are  formed  by  driving  entries  and 
cross  entries  so  as  to  intersect  each  other  at  regular  intervals  of,  usually, 
about  100  yd.  The  entire  field  is  thus  ultimately  divided  into  separate 
squares  or  panels,  each  of  which  has  practically  its  own  system  of  ven- 
tilation. Each  alternate  haulway  may  be  made  an  intake  to  supply  air 
to  one  tier  of  panels,  while  the  next  succeeding  passageway  may  be  used 
as  the  return  to  conduct  the  air  from  each  panel  to  the  foot  of  the  upcast. 
If  more  air  than  usual  is  needed  in  any  one  panel,  it  can  be  obtained  at 
once  by  enlarging  the  opening  in  the  regulator  which  controls  the  air 
for  that  section.  In  case  of  an  explosion  in  any  one  panel,  it  is  not 
usually  communicated  to  the  other  panels.  Extractions  can  be  com- 
menced as  soon  as  a  panel  is  formed;  and  usually  consist  in  driving  a 
heading  across  the  panel  and  opening  the  coal  by  single  or  double  rooms 
or  stalls.  Next,  the  room  pillars  are  carefully  drawn  and  the  roof  inside 
the  peripheral  pillar  of  the  panel  is  allowed  to  fall.  A  high  extraction 
of  coal  can  thus  be  safely  secured  with  a  small  loss  of  timber. 

EXAMPLE  55. — NELMS'  RETREATING  SYSTEM 

Room  and  Pillar  Layout  for  Flat  Coal  Seams. — This  method  insures 
the  operator  a  greater  amount  of  coal  than  when  the  seam  is  worked 
advancing  on  the  room  and  pillar  system.  Since  mining  men  in  the 
United  States  now  recognize  that  our  supply  of  fuel  is  exhaustible,  it 
certainly  behooves  all  operators  to  mine  every  ton  of  coal  possible. 

In  this  retreating  system,  the  main  entries  are  driven  50-ft.  centers 
with  cross-cuts  every  100  ft.  The  middle  entry,  hi  the  three-entry 
system,  is  used  for  the  haulage  road,  being  also  a  main  intake  airway. 
After  turning  a  pair  of  butt  entries  off  the  main,  the  second  crosscut, 
200  ft.  from  the  last  butt  entry,  should  be  a  45-deg.  chute  for  motor 
haulage.  The  dotted  lines  on  the  main  entry  at  the  bottom  of  the  butts 
show  the  position  of  the  "parting."  The  motor,  hauling  25  1  1/2-ton 
cars  comes  in  the  middle  main  entry,  swinging  its  trip  of  empties  in  the 
chute,  the  motor  running  up  the  straight  where  the  drivers  have  stocked 
their  loaded  coal.  The  motor  can  then  pull  its  loaded  trip  outside  and 
the  drivers  proceed  to  distribute  their  cars,  two  drivers  going  in  each 
butt  entry.  The  drivers  make  two  trips,  while  the  motor  makes  one. 


PILLAR   SYSTEMS    FOR   SEAMS 


99 


The  butt  entries  are  driven  on  a  90-deg.  angle  from  the  main  entries, 
and  at  a  distance  of  1400  ft.,  they  intersect  a  set  of  three-face  entries 
running  parallel  to  the  main  entries.  The  butts  are  driven  50-ft.  centers, 
with  crosscuts  every  100  ft.  This  system  of  turning  butts  off  the  mains 
is  an  ideal  one  for  haulage  and  ventilation.  Instead  of  driving  rooms 
off  the  butts  beginning  near  the  main 
entry,  the  rooms  are  started  -from  the 
face-entry  side  and  all  coal  is  worked 
toward  the  main  entries. 

Usually  60  ft.  of  solid  coal  is  left  to 
protect  the  face  entries,  and  60  ft.  to 
also  protect  the  mains.  The  rooms 
are  started  four  at  a  time,  and  as  soon 
as  the  first  four  have  been  driven  50 
ft.,  the  next  four  are  started  on  both 
butts.  The  rooms  are  all  driven  on 
sights  90  deg.  off  the  butt  entry  and 
driven  25  ft.  wide  for  a  distance  of 
240  ft.,  there  being  a  15-ft.  pillar  left 
in  each  room.  The  crosscuts  in  the 
rooms  are  from  80  to  100  ft.  apart,  and 
should  be  " staggered"  across  the 
different  rooms  so  as  not  to  make  a 
weak  place  in  the  roof  by  having  the 
breaks  all  opposite. 


FOUR  PAIRS  OF  BUTT  ENTRIES  WILL 
PRODUCE  1200  TONS  OF  COAL  DAILY 


After  driving  the  rooms  the  full 
distance,  they  should  be  cut  over  to 
the  next  room  by  the  mining  ma- 
chine, the  cut  being  20  ft.  wide.  The 
great  advantage  to  be  gained  in  this 
system  is  the  method  of  not  having 
work  scattered  all  over  a  mining 
territory.  Four  pairs  of  butt  entries,  thus  mined,  will  produce  1200 
cars  of  coal  each  working  day. 

In  Fig.  134,  there  are  32  "machine"  rooms  (Nos.  13  to  28  on  each 


FIG.  134. — Plan  of  Nelms'  Retreating  system. 


side)  working  on  the  pair  of  butts,  and  requiring  32  men  (loaders),  two 
men  having  two  rooms  and  working  them  together.  It  is  the  general 
practice  to  clean  up  one  room  at  a  time  and  so  always  have  coal  to  load 
in  one  room  or  the  other.  Each  machine  loader  receives  six  cars,  thereby 
producing  192  cars  per  day. 


100  MINING    WITHOUT    TIMBER 

There  are  10  pillars  being  robbed  (Nos.  6  to  10  on  each  side),  and 
these  produce  30  cars  of  coal,  as  the  pillars  are  worked  by  one  man.  In 
some  places  two  men  work  the  pillars.  The  "turn"  in  coal  mines  is 
such  that  a  machine  loader  receives  two  cars  to  the  pick  miner's  one, 
thereby  equalling  each  other's  wages,  as  pick  costs  about  twice  as  much 
as  machine  coal.  The  chain  pillar  and  stump  will  produce  12  cars, 
with  four  men  working,  and  the  two  butts  yield  234  cars  per  day. 

The  engineer  can  advance  the  work  in  such  a  standard  way  that  his 
machine  coal  will  always  total  to  the  proper  amount.  A  mine  foreman 
should  find  this  an  easy  way  to  keep  his  men  standardized,  the  machine 
loaders  always  having  machine  places  and  the  pick  men  pick  places, 
thereby  increasing  the  safety  factor  of  his  mine,  as  his  machine  men 
would  never  have  to  do  pick  work. 

The  ventilation  shown  by  the  arrow  heads  is  the  most  practical  to 
use;  the  splits  are  shown  and  also  the  overcast  at  the  bottom  of  the  butt 
entry,  there  being  a  regulator  in  this  overcast.  The  motor  road  is  clear 
of  doors  on  the  main  entries.  The  arrangement  of  chutes  on  the  left 
side  would  be  slightly  different. 

EXAMPLE  56. — NELMS'  ADVANCING-RETREATING  SYSTEM 

Room  and  Pillar  Layout  for  Flat  Coal  Seams. — The  layout  for  the 
advancing-retreating  system  is  as  follows  in  Fig.  135 :  Three  face  entries, 
on  50-ft.  centers,  are  driven  parallel  to  the  main  entries  at  1400-ft. 
intervals.  The  sectional  area  of  the  face  entries  is  kept  as  nearly  as 
possible  to  a  60-ft.  standard  in  a  5-ft.  seam. 

It  is  advisable  where  possible  to  use  all  three  entries  for  intake  air- 
ways; then  the  middle  entry  can  be  used  for  a  haulage  road  and  should 
be  confined  to  itself  and  not  enter  into  the  ventilation  at  all.  No.  1 
room  on  the  butt  entry  can  be  driven  16  ft.  wide  and  used  as  a  return 
airway.  When  the  No.  1  room  is  maintained  for  an  airway,  it  should  be 
widened  toward  the  face  or  main  entry  and  be  driven  on  70-ft.  centers 
with  the  face  entry.  A  30-ft.  pillar  of  solid  coal  should  be  left  between 
No.  1  and  No.  2  rooms;  No.  1  rib  can  then  be  easily  extracted  when 
robbing  is  commenced  on  this  butt  entry. 

The  gob  in  No.  I  room  should  be  kept  as  low  as  possible  and  if  easily 
handled,  it  should  all  be  loaded  and  dumped  outside;  the  result  is  a 
return  airway,  16x6  ft.  =96  sq.  ft.,  which  is  usually  large  enough. 

Butt  entries  should  be  turned  every  450  ft.  A  chute  is  driven  on  a 
60-deg.  angle,  from  the  middle  main  entry  to  the  outside  main  for  haulage. 
The  butts  are  turned  on  a  90-deg.  angle,  and  are  driven  on  50-ft.  centers 
for  a  distance  of  1400  ft.  A  60-deg.  chute  should  connect  the  butt 
entries  at  25  ft.  from  the  center  of  the  outside  main,  for  haulage  from 
the  butt  entry.  It  is  bad  practice  cutting  corners  off  break-throughs. 


PILLAR   SYSTEMS    FOR    SEAMS 


101 


Each  butt  entry  should  maintain  a  sectional  area  of  about  50  sq.  ft. 
and  be  driven  perfectly  straight  so  as  to  overcome  the  troubles  of  track 
laying,  cars  jumping  track,  etc.  Entry  sights  should  never  be  more  than 
180  ft.  apart,  sq  as  to  allow  the  mine  foreman  a  good  chance  to  keep  his 
sights  well  up.  In  providing  -ventilation,  two  pairs  of  butts  on  one  split 
are  suitable.  The  rooms  should  be  turned  90  deg.  off  butts  and  driven 
as  shown  in  plan. 


^-Advancing  Butt  Entry 


Betreuting  Butt  Entry 


=  3 


p 

Solid    Coal 


LL 


Faces  of  Rooms  to  be  all  liven  oil  210'Liuu 
Fid.  135. — Plan  of  Nelms'  Advancing-retreating  system. 


GENERAL  LAYOUT 

No.  1  room  should  be  started  as  soon  as  possible,  then  No.  2  and  so  on 
from  the  advancing  butt  entry.  When  No.  2  room  is  finished  working 
on  the  face,  No.  14  should  just  be  started;  when  No.  2  rib  is  out,  No.  14 
room  should  be  just  finished  and  No.  27  room  just  starting.  The  ribs 
must  be  extracted  as  soon  as  each  room  is  finished,  no  matter  whether 
the  next  room  on  the  advance  side  is  finished  or  not;  the  rib  is  started 
while  the  next  room  is  still  30  ft.  from  being  finished.  When  No.  32  is 
finished  working  on  the  face,  No.  19  rib  is  just  finished,  also  No.  1  room 
on  the  retreating  butt  is  about  finished  and  No.  13  room  just  started. 
As  soon  as  No.  2  rib  on  the  retreating  entry  is  finished,  it  is  advisable  to 
start  extracting  immediately  the  butt-entry  stumps  and  chain  pillar, 
bringing  everything  along  with  the  retreating  butt  and  closing  entry  in 
tight,  knocking  out  the  brattice  in  each  succeeding  break-through  for 
ventilation. 


102 


MINING    WITHOUT   TIMBER 


This  method  allows  the  rib  men  and  the  machine  loaders  to  be  always 
separate,  the  workings  are  confined  to  the  smallest  space  possible  for  a 
large  tonnage,  and  ventilation  is  easy.  The  mine  will  not  be  dotted 
with  old  abandoned  workings  if  the  method  is  consistently  maintained. 

When  a  set  of  butts  are  thus  worked  out  they  are  off  the  operator 's 
hands.  There  is  never  any  danger  of  a  squeeze  as  every  movement  of 
the  rock  runs  up  against  solid  coal,  and  for  this  reason  it  is  impossible  to 
have  a  squeeze  swing  across  a  set  of  butts  to  another  set,  as  it  generally  does, 
where  both  entries  are  worked  advancing.  Five  pairs  of  butts  developed 
on  this  plan  can  produce  from  1000  to  1500  cars  of  coal  a  day. 

.  Small  mule  "partings"  should  be  at  the  bottom  of  each  pair  of  butt 
entries  and  the  distance  for  mule  haulage  will  then  be  at  a  minimum. 
The  coal  from  these  partings  can  be  gathered  by  a  6-  or  8-ton  locomotive, 
and  delivered  to  a  longer  parting  whence  a  larger  locomotive  can  take  it 
outside. 


Boundary 


Barrier  Pillar 


Fio.  136. — Plan  for  pillar-drawing,  Connellsville. 

EXAMPLE  57. — CONNELLSVILLE  COKE  DISTRICT,  WESTERN  PA. 
(See  also  Example  58.) 

Retreating  System  on  Thick,  Flat  Pittsburg  Seam. — The  plan  of  mining 
in  the  Connellsville  region  has  grown  from  the  primitive  methods,  suitable 
to  the  favorable  mining  conditions  when  operations  were  first  started,  to 
the  scientific  methods  which  became  necessary  as  the  cover  increased, 
and  when  most  of  the  difficulties  that  are  likely  to  be  met  in  deep  mining 
were  encountered  and  overcome.  The  conditions  here  are  that  all  the 
coal  is  coked  so  that  fine  coal  is  an  advantage;  no  machines  are  used,  but 
the  coal  is  dug  with  pick;  the  seam  averages  71/2  ft.  of  clean  coal  mined; 
the  roof  is  friable  and  some  coal  is  left  in  the  top  to  support  it;  the  over- 
lying stratum  is  generally  6  to  10  ft.  of  slaty  coal  with  sandstone  above. 

Where  the  acreage  owned  or  assigned  to  each  mine  is  too  large  to 


PILLAR    SYSTEMS    FOR    SEAMS 


103 


admit  of  going  to  the  extreme  boundary  before  starting  to  draw  the 
ribs,  it  is  customary  to  divide  the  field  into  panels  1000  or  1500  ft.  square, 
by  face  headings  driven  off  the  main  butt  headings  as  shown  in  Fig.  136. 
The  coal  is  first  removed  from  the  extreme  side  of  each  panel  away  from 
the  main  butt  headings,  a  diagonal  break  line  is  established,  as  shown, 
and  the  coal  withdrawn,  retreating  toward  the  near  corner,  keeping  the 
break  line  straight,  and  the  coal  between  where  the  drawing  is  being 
carried  on  and  the  main  butt  headings  as  nearly  solid  as  possible,  the 
butt  headings  and  rooms  being  drawn  only  fast  enough  to  open  up  the 
coal  for  the  drawings  The  break  line  of  roof  is  kept  as  nearly  perpen- 
dicular as  practical  to  direction  of  rooms.  The  recent  tendency  in  the 
large  mines  is  not  to  start  pillar-drawing  till  the  boundary  is  reached. 


Fia.  137.— Pillar-drawing  with  thin  cover,  ConnellsvUle. 

The  face  of  the  coal  in  this  region  is  well  defined  on  a  line  running 
N  17°  E.  Where  the  grades  are  not  too  great  all  headings  are  driven 
square  on  the  face  or  on  the  butt,  and  the  rooms  always  on  the  face  and 
only  to  the  rise.  The  rooms  are  driven  10  or  12  ft.  wide  and  a  line  of 
posts  set  as  the  room  advances,  as  shown  in  Fig.  137,  the  posts  being  set 
about  4  or  5  ft.  apart.  A  track  of  wooden  or  steel  rails  is  laid  by  the 
miner  close  to  the  upper  rib.  The  width  of  the  pillar,  which  varies  from 
10  to  70  ft.,  is  governed  by  the  softness  of  the  bottom  and  the  thickness  of 
the  overlying  strata. 

There  is  some  variation  in  the  method  of  drawing  the  individual 
ribs  but  the  principle  is  the  same.  On  account  of  the  nature  of  the 
roof,  short  falls  are  necessary,  two  or  three  being  made  before  the  over- 
lying rock  is  broken.  When  the  rock  breaks  it  will  crush  posts  so  that  it 


104 


MINING    WITHOUT   TIMBER 


is  necessary  to  break  the  roof  against  the  end  of  the  ribs  and  not  over  posts. 
If  care  is  taken  the  digger  knows  when  to  expect  a  fall  and  very  few 
posts  need  be  lost. 

Fig.  137  shows  the  method  in  use  when  the  overlying  strata  are  under 
200  ft.  thick  and  where  a  curved  track  and  track  along  the  face  are  not 
put  in.  A  slab  a  is  taken  off  the  right-hand  rib,  the  whole  of  the  left- 
hand  rib  is  taken  out  excepting  enough  only  left  in  to  keep  the  gob  from 
mixing  with  the  coal  as  shown  at  6.  A  small  shoulder  c  is  left  in  the  far 
corner  of  the  rib  to  take  the  brunt  of  the  weight.  A  row  of  posts  /  is  set 
from  this  shoulder  across  the  room  to  preserve  a  working  face.  The 
posts  e  in  the  back  of  this  row  and  between  it  and  the  previous  fall  or 
break  are  drawn  and  any  which  cannot  be  drawn  are  cut  so  as  to  get  a 


FIQ.  138. — Pillar-drawing  under  thick  cover,  Connellsville. 

good  fall.  Two  men  work  in  each  room.  Room  40  shows  the  situation 
in  the  last  room  of  a  tier  of  rooms  along  a  butt  heading.  Slabs  are  being 
taken  off  each  side  pillar  while  the  men  are  protected  by  the  row  of  posts  d. 

Room  39  shows  the  condition  of  a  room  just  before  producing  a  fall 
by  withdrawing  the  row  of  posts  d  and  separate  posts  e  between  the  rows 
d  and /.  The  row  of  posts/  and  the  stump  of  coal  c  protect  the  face  dur- 
ing the  fall.  The  stump  c  is  next  removed  excepting  for  the  small  slab 
6  and  the  work  proceeds  as  shown  at  the  face  of  room  40. 

Fig.  138  shows  a  method  of  drawing  pillars  where  there  is  over  200  ft. 
of  surface  above  the  coal  and  where  a  curve  is  used  to  run  the  track  along 
the  face.  The  rooms  are  driven  12  ft.  wide  while  the  pillars  between 
the  rooms  vary  from  34  ft.  up  to  60  ft.  All  of  the  rooms  are  driven  on 
sights  so  that  the  pillars  may  be  of  uniform  thickness.  After  the  last 
room  on  the  heading  is  driven  the  required  length,  which  is  about  300  ft., 
the  pillar  is  cut  across  at  the  face  of  the  room  and  20  or  30  ft.  removed 


PILLAR    SYSTEMS    FOR    SEAMS 


105 


before  drawing  the  posts  and  getting  a  fall  of  the  roof.  The  usual 
method  is  then  for  two  men  to  work  on  each  pillar,  while  one  man  cuts 
back  in  the  center  of  the  pillar  on  the  face  of  the  coal  as  far  as  he  can 
conveniently  shovel  as  shown  at  a,  room  27,  the  mine  car  being  on  track 
b,  the  other  man  is  drawing  stump  c  and  shoveling  into  the  same  car. 

When  c  is  all  removed  a  fall  is  made  and  the  situation  is  similar  to 
that  shown  in  rooms  30  and  26,  the  curved  track  having  been  removed 
to  one  side  and  a  straight  one  substituted.  Now  one  man  cuts  into  the 
side  of  the  pillar  8  ft.  from  the  end  at  d  as  shown  while  the  other  is  re- 
moving the  stump  e.  When  this  is  accomplished  a  fall  is  made  and  the 
curve  put  in,  the  conditions  being  then  as  shown  in  room  29;  the  two 
men  then  continue  to  cut  over  toward  the  gob  in  the  next  room  as  shown 
in  room  29,  the  curved  track  having  meanwhile  been  put  along  the  face. 
Room  28  shows  the  situation  when  the  two  men  have  driven  this  cut 
through  to  the  gob  in  the  adjoining  room.  Room  27  illustrates  the  next 


Rib  cut  across  Rib  ready  for  first  fall.       Rib  ready  for  last  fall. 

The  Bnginetnng  $  Mining  Journal 

Fio.  139. — Second  method — pillar-drawing  under  thick  cover,  Connellsville. 

step.  While  one  man  works  on  the  pillar  c  in  the  far  corner  of  the  room 
the  other  starts  the  cut  a  back  into  the  face  as  shown.  The  curved  track 
b  is  then  lifted  and  a  straight  track  put  in  as  shown  in  room  26  in  order  to 
get  out  the  pillar  e.  While  one  man  is  removing  this  pillar  the  other 
one  starts  to  cut  into  the  rib  as  shown  at  d,  room  26. 

When  a  fall  is  to  be  made  posts  are  set  18  in.  apart,  as  shown  in  room 
26  across  the  end  of  the  room  and  along  the  end  of  the  hole  into  the 
pillar.  All  of  the  other  posts  beyond  the  break  rows  are  drawn.  The 
curved  track  is  laid  into  the  new  cut  in  the  side  of  the  pillar  at  d  and  by 
the  next  morning  the  roof  has  fallen.  In  places  where  the  rib  is  wider 
than  shown  on  the  plan  a  couple  of  falls  can  be  made  in  the  width  of  the 
pillars  by  placing  break  rows  of  props  similar  to  those  already  described. 

A  third  pillar-robbing  method,  for  a  heavy  cover,  is  shown  in  Fig.  139. 
It  cross-cuts  the  pillar  for  a  track  and  uses  the  same  curve  as  the  previous 
method,  but  instead  of  cutting  up  the  resulting  pillar-slab  into  blocks, 
it  begins  at  the  gob  and  withdraws  the  slab  gradually  roomward,  mean- 
while recovering  most  of  the  props  in  the  figure  that  were  put  in  to 
protect  the  excavation  of  coal.  A  ratchet-chain  puller  for  props  is  used 
where  necessary.  The  use  of  this  method  at  the  Continental  No.  1  mine 


106  MINING    WITHOUT   TIMBER 

of  the  Frick  Coal  Company  gives  a  recovery  of  about  90  per  cent,  of  the 
coal.  The  losses  arise  from  a  6-in.  coal  layer,  impure  with  sulphur,  left 
on  the  floor;  a  coal  layer,  of  4  in.  in  the  rooms  and  9  in.  in  the  entries, 
left  on  the  roof;  and  some  occasional  stumps  lost  in  robbing  the  pillars. 

The  foremen  of  the  district  are  guided  by  the  following  10  rules  in 
extracting  pillars.  (1)  Pillar  robbing  must  not  be  stopped  or  diverted 
from  the  line  of  fracture  without  the  consent  of  the  chief  engineer.  (2) 
Robbing  must  proceed  from  the  new  toward  the  older  gob  to  prevent 
uncalculable  pressure  on  the  working  face.  (3)  Ribs  must  be  robbed 
within  one  month  of  driving  rooms.  (4)  Room  centers  must  be  at  the 
prescribed  distance  apart.  (5)  In  robbing  entry-pillars,  a  length  of 
only  2  room-widths  must  be  attacked  at  once  along  the  line  of  fracture. 
(6)  Water  ditches  must  be  made  for  entry-drainage  and  especial  care 
must  be  taken  on  soft  bottoms.  (7)  A  200-ft.  pillar  must  be  left  on 
each  side  of  the  main  or  flat  entry  during  its  life.  (8)  A  wide  barrier 
pillar  must  be  left  and  care  must  be  taken  in  approaching  a  neighbor's 
boundary.  (9)  Before  permitting  a  fall  of  the  roof,  all  timber  must  be 
drawn  and  a  passage  left  for  the  escape  of  the  miners.  (10)  A  miner 
should  keep  the  pillar  he  is  drawing  between  himself  and  the  gob 
instead  of  working  between  gob  and  pillar. 

EXAMPLE  58. — PITTSBURG  BITUMINOUS  DISTRICT,  WESTERN  PA. 
(See  also  Example  57.) 

Advancing-retreating  or  Retreating  System  in  Panels  on  Thick  Flat 
Pittsburg  Seam. — In  those  mines  of  western  Pennsylvania,  extracting 
the  thick  Pittsburg  coal  seam  for  shipment  to  market,  the  mining  layout 
is  different  from  that  of  the  coking  district  of  Example  57.  Since  the 
policy  of  the  market-coal  mines  is  to  obtain  as  much  lump  coal  as  possible, 
the  bulk  of  the  coal  is  obtained  from  the  rooms,  for  pillar  coal  is  bound  to 
be  more  or  less  crushed.  This  policy  requires  wide  rooms  and  narrow 
pillars  and  results  in  a  lesser  total  recovery  of  coal,  but  as  an  offset,  more 
coal  can  be  won  by  machine  cutters  which  work  advantageously  in  this 
thick  seam.  By  the  deep  and  fast  undercutting  possible  with  machines, 
blasting,  with  its  ensuing  slack,  is  at  a  minimum;  and  progress  is  rapid 
enough  to  preserve  the  coal  faces  from  long  exposure  to  the  atmosphere 
and  to  allow  of  systematic  timbering  and  an  even  subsidence  of  the  roof. 

The  Monongahela  River  Cons.  Coal  and  Coke  Company  is  a  very 
large  producer,  and  a  composite  drawing  of  its  method  of  working  is 
shown  in  Fig.  140.  Above  the  upper  "butts"  or  butt  entries,  the  rooms 
have  not  advanced  far  from  the  "Face  Entries."  On  the  middle  butts, 
the  rooms  have  reached  the  end  of  the  upper  panel  and  pillar-drawing 
has  advanced  halfway.  On  the  lower  butts,  the  lower  panel  is  being 
worked  by  retreating  from  the  panel-end,  the  rooms  are  nearly  completed, 


PILLAR    SYSTEMS    FOR    SEAMS 


107 


and  the  line  of  roof-fracture,  across  both  upper  and  lower  panels,  is 
following  the  pillar-drawing  and  is  not  far  behind  the  finished  rooms. 

In  the  advancing  system,  the  panels  off  the  upper  butt  entry  would 
be  attacked  first,  and  the  rooms  of  this  advancing  panel  would  end  in  the 
old  gob,  while  the  rooms  of  the  lower  or  retreating  panel  would  end 
against  the  solid.  As  shown  in  the  figure,  the  room-centers  are  39  ft. 


three  Fairs  of  Butts  on  Slip 

Intake  Current  > 

Return  Curren 
Brick  Stoppings    = 
Doors  /"• 

Overcasts 


Fia.  140. — Fkst  layout  (Monongahela  colleries)  at  Pitteburg,  Pa. 

apart,  of  which  space  the  pillars  occupy  15  ft.  It  will  be  noticed  that  the 
room-stumps  of  each  upper  panel  are  left  undisturbed  on  the  advance  so 
as  to  protect  the  return  airway,  but  when  the  pillars  of  the  lower  panel 
are  being  drawn,  the  upper  stumps  are  also  pulled,  as  the  receding  line  of 
roof-fracture  passes  them,  along  with  the  butt-entry  pillars.  The  last 
pillars,  however,  must  be  left  undisturbed  in  the  advancing  system  along 


108 


MINING    WITHOUT    TIMBER 


their  whole  length  until  all  the  adjoining  coal  has  been  exhausted  up  to 
the  boundary.  The  use  of  four  main  entries,  by  this  method,  allows  the 
two  outside  gangways  to  be  return-airways  and  the  two  intake-airways 
to  be  on  the  inside  and  thus,  gives  an  ample  main-airway  area  and  a  min- 
imum interference  with  transport.  The  room-work  is  in  the  fresh  air  and 
pillar-drawing  is  on  the  return-air  side  of  it.  The  room-track  is  always 
laid  along  the  straight  rib,  and  in  many  mines  the  refuse  between  the 
track  and  the  other  rib  fills  the  room  nearly  roof  high. 


FIG.  141.— Second  layout  for  large  output,  Pittsburg. 


Fig.  141  shows  a  second  layout  for  large  output,  used  in  the  Pittsburg 
seam,  with  six  main  entries.  There  are  three  face  entries,  nominally,  but 
four  actually,  as  the  nearest  room  on  the  butt  is  advanced  along  with 
them  so  as  to  give  an  additional  airway.  As  shown,  the  rooms  are  only 
worked  on  the  outbye  side  of  the  butts,  and  the  first  room  is  started 
from  the  far  end  of  a  panel  and  followed,  at  the  proper  distance  on  the 
retreat,  by  pillar-drawing.  By  starting  work  from  No.  2  and  the  follow- 
ing butts  at  the  proper  time,  it  is  possible  to  keep  the  line  of  roof-fracture 


PILLAR   SYSTEMS    FOR   SEAMS 


109 


of  a  panel  continuous,  for  entry-pillars  and  room-stumps  are  removed 
as  shown  from  the  panel-end  back  to  the  butts. 

The  method  of  Fig.  141  has  permitted  the  extraction  of  70  per  cent, 
of  the  pillars  by  machine  cutters  under  an  average  cover  of  200-ft.  thick- 
ness. For  this  purpose  machine  cross-cuts,  21  ft.  wide,  are  made  in  the 
pillars  so  as  to  leave  for  each  a  stump  only  9  ft.  wide  to  be  removed  by 
hand-pick.  This  cross-cutting  is  shown  by  the  different  cross-hatching 
of  the  figure  which  also  illustrates  the  overcasts  and  brattices  for  ventila- 
tion, and  the  chutes,  etc.,  for  transport. 

A  third  method  of  attack  by  which  one  company  mines  over  2,500,000 
tons  yearly  is  shown  in  Fig.  142.  Here  the  room  pillars,  after  the  room 


FIG.  142. — Third  layout,  with  tapering  pillars,  Pittsburg. 

has  advanced  100  ft.,  or  to  the  first  break-through,  are  gradually  tapered 
off  to  a  point  at,  the  room-end.  This  causes  the  roof  to  fall  along  the 
tapered  parts  of  the  pillars  and  the  latter  are  lost,  but  much  of  the  thicker 
pillar  near  the  room-neck  can  be  recovered  by  subsequent  careful  pick- 
work.  This  method  gets  nearly  all  the  coal  by  room-work,  and  a  total 
recovery  of  90  per  cent,  is  claimed  by  its  advocates.  It  is  more  danger- 
ous, however,  than  the  two  previous  systems,  requires  more  timber,  and 
squeezes  are  more  liable  to  occur. 

Where  this  high  Pittsburg  seam  is  dirty,  so  that  much  gob  must  be 
stowed  along  the  rib  on  the  advance,  it  is  customary  on  drawing  the 
pillars  to  leave  a  vertical  shell  of  from  12  to  18  in.  of  coal  next  to  the 
gob  to  prevent  any  pollution  of  the  broken  coal. 


CHAPTER  XXI 

FLUSHING  SYSTEM  FOR  FILLING  SEAMS  AND  RECOVERING 

PILLARS 

EXAMPLE  59.— ANTHRACITE  DISTRICT,  EASTERN  PA. 
(See  also  Examples  5,  51  and  59.) 

Parallel  Seams  of  Various  Thickness  and  Dip  Filled  with  Refuse  from 
Breakers  and  Dumps. — The  flushing  system  was  first  developed  in  1885 
at  the  Pardee  No.  5  mine  near  Hazleton,  Pa.,  and  was  later  copied  and 
extensively  used  in  many  German  collieries.  Three  conditions  made 
flushing  a  valuable  innovation  in  the  Pennsylvania  anthracite  region, 
namely,  the  numerous  large  dumps  of  waste  available  for  filling,  the 
parallel  and  superincumbent  seams  to  be  extracted,  and  the  overlay  of 
much  workable  coal  by  townsites.  The  gravity  of  the  urban  situation 
is  evidenced  by  the  report  of  April,  .1911,  made  by  the  Scranton  Com- 
mission. l  This  report  states  that  a  large  part  of  Scranton  is  already 
undermined  and  that  for  the  stability  of  the  present  dangerous  area  of 
15  per  cent,  of  the  city  and  for  the  recovery  of  the  coal  pillars  from  the 
balance  the  flushing  system  is  the  only  remedy. 

The  following  description  in  based  on  the  author's  visits  to  mines  of 
the  following  coal  companies:  Philadelphia  and  Reading;  Delaware 
and  Hudson;  Delaware,  Lackawanna  and  Western;  Plymouth;  and 
Lehigh  Valley. 

In  mining  the  flat  seams  to  the  north  of  Wilkesbarre  by  the  pillar 
system  of  Fig.  130  much  of  the  waste  broken  with  the  coal  can  be  left 
in  the  rooms;  but  in  the  seams  of  the  southern  districts  where  mining  is 
done  by  "overhand  stoping  with  shrinkage  and  chutes,"  as  in  Figs.  132 
and  133,  all  the  waste  has  to  be  hoisted.  The  crude  coal  reaching  the 
surface  is  a  mixture  of  pure  coal,  "slate,"  "slate-coal",  and" bone." 
The  "slate"  corresponds  to  the  shale  and  clay  of  the  partings  and  beds 
of  the  bituminous  regions,  the  "slate-coal"  consists  of  lumps,  part  pure 
coal  and  part  slate,  and  the  "bone"  is  a  coal  containing  too  little  carbon 
(present  limit  60  per  cent.)  to  be  marketable.  All  crude  coal  is  put 
through  a  dressing  mill  or  "  breaker"  in  which  impure  pieces  are  broken 
sufficiently  to  detach  the  slate  and  bone  from  the  pure  coal,  so  that  all  the 
latter  may  be  screened  for  separation  into  commercial  sizes  and  the 
former,  along  with  the  "culm"  or  coal  dust,  be  sent  to  the  waste  dump 
or  mine  stopes. 

*  "Mine  Caves  under  Scranton,"  by  E.  T.  Conner.  Trans.  Min.  Eng.,  Vol.  42,  p.  246. 

110 


FLUSHING    SYSTEMS    FOR    FILLING    SEAMS 


111 


112  MINING    WITHOUT    TIMBER 

The  limit  of  size  between  fine  coal  and  unmarketable  "culm"  has  so 
decreased  in  recent  years  that  now  all  the  fine  sizes  of  the  breakers,  as 
well  as  many  old  waste  dumps,  are  being  washed  over  shaking  screens 
in  special  mills  called  "washeries,"  for  their  content  of  fine  coal  of 
commercial  value.  The  present  upper  limit  for  "culm"  is  a  diameter 
varying  from  5/64  to  3/16  in.,  but  some  independent  operators  use 
also  some  larger  sizes  for  flushing.  This  rejected  fine  coal,  as  mixed  with 
the  slate  and  bone  tailing  from  the  breaker  and  the  ashes  from  the  boiler 
plant,  forms  "  slush, "  the  chief  material  now  used  in  filling  the  mine  stopes 
by  the  flushing  system.  At  some  mines,  the  larger  pieces  of  bone  are 
saved  on  a  special  dump  as  of  possible  future  value.  Fig.  143  shows 
the  Dodson  colliery  at  Plymouth,  Pa.,  with  the  waste  dump  at  A,  the 
breaker  at  B,  and  the  washery  at  C. 

In  digging  an  old  waste  dump  for  passage  through  a  washery,  in 
order  to  separate  the  marketable  coal  before  flushing,  a  system  of  chain 
conveyors  as  at  D  and  H,  Fig.  143,  is  used.  The  usual  conveyor  has  a 
single  chain  and  drags  its  steel  plate  scrapers,  18  in.  long,  12  in.  high, 
and  set  3  ft.  apart,  in  a  trapezoidal  trough  made  by  lapping  the  ends  of 
3-ft.  lengths  of  steel  or  cast-iron  plate.  The  maximum  length  of  a  single 
conveyor  trough  is  about  500  ft.  It  is  supported  within  square  wooden 
frames,  E,  set  8  ft.  apart,  and  built  of  4x6-in.  pieces.  Near  the  top  of 
the  frames  E,  run  two  25-lb.  steel  rails  to  support  the  scrapers  on  their 
return  trip. 

Each  conveyor  is  run  by  an  independent  steam  engine,  as  at  F,  con- 
nected by  gearing  to  its  head  end,  and  its  capacity  of  100  to  200  tons  of 
dry  material  per  hour  is  fed  in,  anywhere  along  the  trough,  by  hand 
shovels  or  by  hydraulicing  with  hose.  Obstacles  between  the  dump 
and  washery  are  passed  by  using  several  conveyors,  set  at  an  angle,  of 
which  only  the  conveyor  at  the  feed  end  need  be  shifted  as  the  dump 
dwindles.  The  driving  engine  is  set  on  a  timber  frame  so  that  it  can  be 
easily  pushed  into  line,  by  screw  jacks,  when  the  conveyor  is  moved 
over  by  levers;  both  engine  and  conveyor  are  elevated  on  rollers  before 
shifting.  This  is  done  by  the  regular  attendants  who  consist  of  two  men 
for  feeding  and  one  man  at  each  driving  engine. 

So  much  fine  marketable  coal  can  now  be  saved  from  the  present 
breaker-tailing  and  from  the  old  "culm"  dumps  that  the  final  rejected 
waste  can  fill  only  a  fraction  of  the  space  left  above  the  gob  in  the  under- 
ground rooms.  The  huge  dumps  which  formed  such  a  prominent 
feature  of  the  landscape,  as  late  as  the  early  nineties,  are  rapidly  dis- 
appearing and  filling  is  now  being  won  even  from  river  beds. 

Thus  the  Plymouth  Coal  Co.  has  a  plant  to  bring  sand  and  fine  mine 
waste,  now  settled  at  the  bottom  of  the  Susquehanna  river,  to  its  Dodson 
mine  No.  12.  A  suction  pump  on  a  barge  located  across  the  river  from 
the  Dodson  breaker  delivers  into  a  pipe  which  crosses  the  river  on 


FLUSHING    SYSTEMS    FOR   FILLING    SEAMS  113 

barges  and  discharges  into  an  elevator  which  lifts  the  material  to  the 
flushing  flume  for  the  mine. 

As  no  pieces  larger  than  1-in.  dia.  and  few  over  1/4-in.  dia.  are  used 
in  flushing,  the  coarser  pieces  of  bone  and  slate,  from  breaker  or  dump, 
are  passed  through  a  pulverizer  usually  of  the  Williams'  or  Jeffrey's  type, 
before  reaching  the  flume  G,  Fig.  143,  where  they  mix  with  the  fine 
waste  from  washery  C.  Enough  water  is  put  in  the  flume  to  transport 
the  slush  along  the  flat  pipes  above  the  stopes,  so  the  liquid  slush  carries 
only  about  20  per  cent,  of  solids  by  weight.  The  descent  of  the  slush  is 
through  wrought-iron  pipes,  4  to  6  in.  dia.,  following  either  a  shaft  or  a 
special  bore-hole  into  the  workings,  perhaps  1000  ft.  beneath.  Over  the 
top  of  the  descending  pipe  is  placed  a  funnel  and  a  screen  with  1-in.  holes, 
and  at  each  flushing  station  are  three  gate  valves,  to  regulate  the  flow 
into  the  stopes,  connected  by  electric  signals  with  the  surface.  One  of 
these  station  valves  regulates  the  flow  horizontally,  another  cuts  off  the 
vertical  column  below,  a*nd  by  a  third  the  column  can  be  drained  up  to 
the  nearest  flowing  point  above,  in  case  of  a  stoppage. 

The  pipes  used  for  transport  along  the  levels  are  4  to  6  in.  dia.,  and  of 
either  wrought-iron  or  wood.  In  upgrade  levels  or  where  the  pipe  is  under 
much  pressure,  new  iron  pipes  with  screw  or  flange  couplings  must  be  used, 
but  when  these  are  somewhat  worn  they  are  transferred  to  the  down- 
grade levels.  In  the  latter,  the  iron  pipe  has  standard  couplings  on 
tangents,  but  on  curves  it  has  7-in.  unthreaded  nipples  for  couplings 
slipped  over  the  pipe  ends  and  made  tight  by  wooden  wedges.  These 
wedged  couplings  enable  the  pipe  to  be  rotated,  when  its  bottom  gets 
thin,  so  that  it  can  be  worn  to  a  mere  shell  all  around  before  rejection. 

The  wooden  pipe  is  made  in  Elmira,  N.  Y.,  of  tenoned  maple  staves 
about  2  1/2  in.  thick  which  are  bound  with  spiral  steel  hoops  and  coated 
with  tar.  It  comes  in  2-  to  8-ft.  lengths,  with  male  and  female  ends  for 
slip-jointing  with  cement.  It  can  be  joined  to  cast-iron  fillings  by  in- 
serting special  cast-iron  nipples  in  its  ends,  and  its  own  joints  can  readily 
be  made  to  follow  easy  curves.  It  is  lighter  and  cheaper  than  iron  pipe, 
is  found  to  last  well  on  downgrade  levels,  and  is  preferable  for  use  with 
acid  water.  In  one  mine,  the  iron  pipe  is  used  on  a  2-mile  permanent 
transportation  line  and  the  wooden  pipe  in  the  neighborhood  of  the 
actual  filling. 

Plugged  cast-iron  tees  are  placed  at  100-ft.  intervals  along  all  lines  in 
the  levels,  so  any  obstruction  can  be  easily  located  and  removed.  When 
a  pipe  is  upgrade,  a  special  precaution  is  taken  against  clogging  by  pass- 
ing fresh  water  alone  through  it,  for  15  min.,  before  stopping  the  flow. 
Care  must  be  taken  to  provide  air  escapes  at  high  points  of  the  lines  in 
order  to  avoid  water  hammer. 

The  openings  filled  by  flushing  are  old  rooms  opened  on  the  pillar 
system  of  the  last  chapter.  A  room  on  a  dip  is  easiest  filled,  as  it  requires 


114 


MINING    WITHOUT   TIMBER 


only  one  dam  or  barrier  at  its  lower  end.  One  disadvantage  of  increasing 
steepness  is  the  greater  strength  of  dam  necessary  to  resist  the  corre- 
spondingly higher  water  head.  In  the  Dorrance  mine  the  old  rooms  had 
been  opened  on  the  rise  from  double  flat  entries  as  in  Fig.  130.  Every 
ten  rooms  along  the  entry  were  separated  by  panel-pillars  following  the 
dip.  For  filling,  the  flushing  pipe  was  laid  along  the  airway  above  the 
rooms  and  its  discharge  placed  at  the  head  of  the  central  room  of  a  panel 
of  empty  rooms.  The  latter  had  been  prepared  for  filling  by  erecting 
dams  across  the  necks  at  the  room-bottoms  and  behind  the  break-through 
brattices  of  the  ninth  room 's  pillar,  for  the  last  room  of  the  panel  was  to  be 
left  open  as  an  air  and  manway.  The  brattices  of  the  break-throughs  of 
the  intermediate  rooms  had  been  removed  to  permit  of  a  free  flow  of 
filling  along  the  panel. 

Room  dams  are  made  of  either  stone  or  wood.     The  former  are  thick 
walls  of  roof  slate  laid  up  with  mortar  of  slush  and  straw  in  a  similar 


Cross  Sec. 


Sec.  a-b 
FIG.  144.— Dam  for  holding  slush,  Eastern  Pa. 

form  to  the  wall  of  Fig.  145  described  in  the  next  Example.  The  favorite 
dams  are  of  wood  and  a  typical  one  is  shown  hi  Fig.  144.  Round  props 
ab,  of  sufficient  size  for  the  expected  strain,  are  covered  on  their  upper 
side  with  2-in.  plank  and  backed,  as  an  extreme  case,  with  stringers  66' 
and  cc'  with  corresponding  angle  braces  &/  and  cd.  If  the  seam-walls  are 
strong,  the  hitches  alone  will  hold  the  props,  so  that  the  pieces  66'  and  bf 
can  be  omitted,  and  in  thin  seams  even  cc'  and  cd  are  left  out. 

When  wetted,  the  seams  between  the  planks  soon  close  up  sufficiently, 
but  the  irregular  spaces  around  the  periphery  mn'b'b  are  caulked  with 
straw  in  one  mine,  and  in  another,  with  a  weak  floor,  the  props  are  set  in 
a  low  concrete  wall,  12  in.  wide.  In  one  seam  of  the  Dorrance  mine  on 
an  18-deg.  slope  with  rooms  300  ft.  long,  the  wooden  dam  of  Fig.  144  is 
strengthened  by  a  dry  wall  of  roof  slate,  3  to  5  ft.  thick,  laid  above  the 
plank  ab.  By  slow  flushing  at  first,  this  dry  wall  gets  packed  solid  and 
keeps  the  plank  from  bulging  under  the  heavy  final  pressure  due  to  a 
vertical  water  head  of  90  ft. 


FLUSHING    SYSTEMS    FOR    FILLING    SEAMS  115 

A  room  in  the  Dodson  mine  in  the  22-ft.  Red  Ash  seam  was  worked 
in  two  slices,  the  first  taking  only  8  ft.  of  coal  from  the  floor.  When 
preparing  for  flushing,  the  upper  14-ft.  slice  of  coal  was  not  taken  down 
over  the  neck  for  24  ft.  from  the  room's  lower  end,  so  that  the  subsequent 
wooden  dam  needed  to  be  only  8  ft.  high.  Holes  are  bored  into  the  plank 
of  the  dams  near  the  top,  if  necessary,  to  let  the  overflow  water  escape, 
but  a  better  arrangement  for  steep  dips  is  a  wooden  drain-launder  bk 
laid  on  the  floor  up  through  the  dam  into  the  room.  The  top  m  of  the 
cover  of  launder  bk  is  kept  a  short  distance  above  the  top  of  the  settled 
slush  at  n  by  adding  new  cover-boards  as  the  filling  rises.  The  overflow 
water  then  runs  over  into  the  launder  at  m  and  descends  into  the  gang- 
way ditch  at  g  to  flow  to  the  sump;  whence  it  is  pumped  to  the  surface, 
where,  being  acid,  it  is  not  reused  unless  fresh  water  is  scarce.  At  the 
Dorrance  mine  where  the  rooms  were  being  filled  on  the  advance  by 
extending  the  flushing  pipe  from  one  panel  of  ten  rooms  to  the  next,  it 
was  the  practice  to  give  each  panel  another  dose  of  slush,  while  with- 
drawing the  pipe,  in  order  to  close  up  the  many  spaces  between  the  settled 
slush  and  the  roof  that  had  developed  since  the  advance.  For  nearly 
flat  seams,  dams  are  built  in  the  openings  all  around  a  panel  of  rooms, 
and  the  end  of  the  flushing  pipe  shifted  around  inside  the  panel,  close  to 
the  roof,  so  as  to  fill  all  portions  equally.  More  or  less  methane  is  given 
off  if  the  slush  is  exposed  to  air  currents,  but  these  are  feebler,  the  smaller 
the  spaces  left  between  slush  and  roof.  As  another  safeguard  against 
gas,  the  filled  panels  are  connected  with  the  return  airways  of  the 
active  mine. 

In  the  considerable  areas  where  a  subsidence  of  the  surface  is  imma- 
terial, the  anthracite  seams  are  best  worked  to  the  boundary,  by  that 
pillar  system  of  the  last  chapter  most  appropriate  to  the  given  conditions; 
and  the  pillars  then  recovered  on  the  retreat,  allowing  the  roof  to  fall. 
Under  the  river  flats  where  roof-falls  might  cause  a  crack  up  to  the  surface 
and  flood  the  workings,  one  company's  mines  are  laid  out  with  permanent 
pillars  of  a  size  just  sufficient  to  sustain  the  roof  indefinitely,  which  means 
16-ft.  pillars  and  24-ft.  rooms  for  depths  of  less  than  400  ft. 

Flushing  as  a  preliminary  to  pillar-drawing  is  beneficial  in  the  anthra- 
cite region  under  two  conditions.  First,  where  the  workings  are  overlaid 
by  virgin  parallel  seams,  and  second,  where  they  are  overlaid  by  townsites. 
Former  market  conditions  made  the  thin  seams  unpayable,  so  that  the 
proper  method  of  exhausting  overlying  coal  seams  from  the  top  down- 
ward was  not  applied.  Now  the  pillars  can  only  be  recovered  from  the 
lower  seams,  without  wrecking  those  above,  by  a  preliminary  filling  of 
the  adjoining  rooms.  Formerly,  it  was  not  thought  that  the  pillars  left 
under  townsites  would  ever  be  worth  recovering,  but  higher  coal  prices 
have  made  them  valuable  and  filling  must  precede  their  recovery. 

The  aforementioned  Scranton  commission  recommends  that  as  slush 


116  MINING    WITHOUT    TIMBER 

alone  has  insufficient  crushing  resistance  for  thick  covers,  sand  should 
be  used  for  filling  under  Scranton  at  depths  beyond  500  ft.  Also  that 
filling  should  begin  in  the  lowest  seam  of  the  series  and  continue  upward 
until  all  are  filled,  care  being  taken  to  have  the  flushed  areas  over  one 
another.  After  all  the  openings  in  all  the  seams  have  been  filled,  the 
pillars  in  the  top  seam  may  be  removed  and  replaced  at  once  by  filling. 
The  next  seam  below  may  not  be  attacked  and  handled  in  like  manner 
until  the  pillars  above,  within  a  large  panel,  are  removed  and  the  over- 
burden has  come  to  rest  on  the  new  filling.  In  this  manner  several 
parallel  seams  could  be  robbed  of  pillars  simultaneously,  by  panels 
retreating  in  vertical  echelon,  the  robbing  in  the  highest  seam  being 
farthest  from,  and  that  in  the  lowest  seam  nearest  to  the  boundary. 

In  some  mines  with  irregular  layouts  and  small  pillars,  the  formation 
had  moved  considerable  before  flushing  was  inaugurated.  Thus  in  the 
22-ft.  Red  Ash  vein  of  the  Dodson  mine  at  Plymouth,  the  overlying 
formation  moved  so  freely  that  gangways  could  only  be  kept  open  by 
using  heavy  timbers  and  brushing  the  floor.  While  in  the  Black  Diamond 
mine  at  Luzerne,  the  walls  of  the  6-ft.  Cooper  seam  were  distorted  with 
frequent  roof-falls,  and  in  the  8-ft.  Bennett  seam  the  roof  had  bent 
enough  to  badly  squeeze  many  of  the  pillars. 

The  seams  of  the  latter  mine,  which  were  excavated  on  the  system  of 
Fig.  130,  dip  about  10  deg.  and  the  pillars  of  the  flushed  portion  are  now 
being  robbed  and  replaced  by  slush.  Where  pillars  are  20  ft.  wide,  or 
more,  an  8-ft.  heading  is  driven  on  one  side  of  the  pillar  on  the  rise,  often 
leaving  a  thin  shell  of  coal  next  to  the  filling.  Then,  when  the  airway 
above  is  reached,  the  balance  of  the  pillar  is  drawn  on  the  retreat.  The 
advance  heading  must  be  well  propped,  but  the  timber  is  mostly  recov- 
ered on  the  retreat  and,  owing  to  the  moving  formation,  the  pillar  coal 
is  so  squeezed  that  but  little  blasting  is  necessary.  Too  much  roof  pres- 
sure sometimes  so  crushes  the  coal  that  it  falls  to  powder  when  extracted. 

The  Dorrance  mine  is  under  a  suburb  of  Wilkesbarre  and  the  policy 
of  the  owner,  the  Lehigh  Coal  Company,  is  to  refrain  from  taking  all  the 
pillar  coal,  when  robbing  flushed  areas  under  cities,  because  an  unsup- 
ported cover  will  settle  down  at  least  10  per  cent,  of  the  coal's  thickness; 
and  with  flat  seams,  where  filling  close  to  the  roof  is  impractical,  the  sub- 
sidence may  be  20  per  cent.  In  fact,  the  surface  in  some  cases  has  sub- 
sided less  from  robbing  pillars  in  open  than  in  filled  seams;  for  in  the 
former  case  local  breaks  of  roof  may  fill  up  the  rooms  with  boulders  and 
support  the  cover,  while  robbing  pillars  completely  in  the  latter  case 
starts  the  whole  cover  to  subsiding  as  in  the  longwall  system. 

The  flushed  workings  observed  in  the  Dorrance  mine  were  on  a  dip 
of  18  deg.  and  on  a  layout  like  Fig.  130  with  rooms  20  ft.  and  pillars  40  ft. 
wide.  A  heading  was  first  driven  up  in  the  pillar,  to  slab  off  18  ft.  of 
coal  alongside  the  filling,  and  on  the  retreat  from  the  room 's  upper  end 


FLUSHING    SYSTEMS    FOR   FILLING    SEAMS  117 

a  24-ft.  cross-cut  was  put  through  the  pillar,  halfway  between  the  original 
12-ft.  break-throughs,  100  ft.  apart.  Thus  after  flushing  the  new  pillar 
openings,  the  roof  was  left  supported  by  a  line  of  coal  pillars  22  feet 
wide  by  32  ft.  along  the  dip. 

Under  Mahonoy  City,  the  22-ft.  Mammoth  seam,  dipping  55  to 
60  deg.,  is  being  worked  in  two  slices  by  a  system  like  that  of  Fig.  131. 
The  lower  slice  of  15  ft.  is  taken  out  in  the  room  on  the  advance  and  the 
upper  7-ft.  slice  allowed  to  fall  into  the  chute,  by  pulling  the  props,  on 
the  retreat.  After  flushing,  the  pillar  is  taken  out,  likewise,  in  two 
slices,  by  driving  a  heading  through  its  center,  leaving  only  a  thin  shell 
of  coal  on  each  side  to  keep  out  the  room-filling.  The  entry  pillars  are 
drawn  on  the  retreat,  and  all  the  new  openings  are  flushed.  In  spite  of 
this  extraction  of  practically  all  the  pillars,  the  surface  here  is  stable,  for 
with  seams  of  steep  dip,  the  subsidence  upon  the  filling  is  not  so  serious 
as  it  is  in  the  case  of  the  flatter  seams  under  Wilkesbarre. 

As  already  mentioned,  the  22-ft.  seam  in  the  Dodson  mine  is  also 
worked  in  two  slices  but  with  the  thin  slice  below.  The  pillars  here  are 
26  ft.  and  the  room  is  24  ft.  wide.  The  lower  slice  of  both  room  and  pillar 
is  mined  on  the  advance  and  the  upper  slice  is  recovered  on  the  retreat 
as  described  in  the  last  paragraph,  except  that  the  layout  follows  Fig.  130 
to  suit  the  12-deg.  dip. 

EXAMPLE  60. — ROBINSON  GOLD  MINE,  RAND  DISTRICT.     TRANSVAAL 
Parallel  Sloping  Beds  Filled  with  Mill  Tailing 

In  spite  of  the  extensive  areas  excavated  since  1885  in  the  conglom- 
erate of  the  Rand,  but  little  filling  has  yet  been  done.  At  a  few  rich 
outcrop  mines,  it  is  true,  the  rooms  were  packed  with  rock  to  enable  the 
pillars  to  be  recovered.  But  packing  is  too  costly  a  method  for  most 
of  the  area.  As  the  mines  reach  depths  exceeding  4000  ft.,  the  former 
sized  pillars  are  proving  too  small,  and  several  unexpected  collapses 
have  occurred.  Recently  the  flushing  system  has  been  tried  with  suc- 
cess at  the  Robinson  mine,  to  permit  of  the  removal  of  some  rich  pillars 
just  under  the  stamp  mill,  in  the  following  manner. 

The  tailing  is  washed  from  the  dump  by  a  1-in.  water  pipe  into  a 
launder,  6  in.  sq.,  which  runs  to  the  top  of  a  winze.  Here  the  pulp  enters 
a  similar  launder  which  descends  along  the  40-  to  50-deg.  dip  of  the  seam 
floor  to  the  ninth  level  of  the  mine.  The  stope  to  be  filled  has  been 
dammed  at  the  lower  end  by  a  dry  wall  W,  see  Fig.  145,  strengthened 
by  poles  P,  and  similar  partition  walls  are  built  at  right  angles  to  cut  it 
up  into  longitudinal  panels.  The  fine  waste  B  is  piled  above  W,  and 
covered  with  old  matting  M  from  the  cyanide  tanks.  When  flushing 
begins,  the  sand  settles  quickly,  the  water  filters  through  the  matting 
and  dams,  whence  it  runs  to  the  sump  to  be  pumped  to  the  surface. 


118 


MINING    WITHOUT   TIMBER 


This  water  is  used  again  after  a  little  lime  has  been  added  to  neu- 
tralize its  acidity  and  render  any  entrained  colloids  harmless  to  hinder  a 
quick  settling.  To  save  water,  the  launders  are  kept  on  a  minimum 
gradient  of  10  deg.  The  water  used  is  6  to  10  per  cent,  of  the  tailing  by 
weight.  The  cost  of  filling  is  given  at  2.1  d  per  ton,  but  as  only  100  tons 
of  tailing  are  sent  underground  daily,  this  probably  does  not  include 
wear  of  the  launders.  The  filling  sets  hard  in  2  or  3  days.  When  a  stope 
is  completely  filled,  it  only  settles  10  per  cent,  of  its  height  when  crushed 
by  the  formation  after  the  pillars  have  been  removed. 

The  residual  cyanide  of  the  tailing  leaving  the  mill  has  been 
destroyed  by  exposure  on  the  old  dumps,  so  that  no  poisonous  results 
have  so  far  ensued  from  using  tailing  as  mine  filling.  In  order  to  utilize 
fresh  tailing,  the  cyanide  must  first  be  rendered  innocuous.  This  is  not 


Fid.  145. — Dam  for  holding  slush,  TransvaaL 

urgent  at  present,  because  the  old  tailing  dumps  are  immense.  Flushing 
the  leading  vats  direct  into  the  mine,  however,  would  save  the  expense 
of  conveying  the  tailing  to  the  top  of  the  very  high  dumps  and  of  redig- 
ging  it  before  flushing.  Hence  some  mines  are  now  getting  ready  for 
direct  flushing. 

The  flushing  system  is  now  being  freely  used  in  the  Rand  to  fill  stopes 
not  under  buildings,  in  order  to  prevent  the  damage  to  the  workings  and 
the  shaft  pillars  which  is  liable  to  ensue  from  pillar-drawing,  especially 
as  the  mines  get  deeper.  The  rock  tailing  available  for  filling  is  much 
more  resistant  to  crushing  than  the  anthracite  refuse  of  Example  59, 
and  is  strong  enough  for  a  filling  at  any  workable  depth.  On  the  central 
Rand,  there  are  two  contiguous  parallel  beds,  the  Main  Reef  below,  and 
the  Main  Reef  Reader  above,  separated  by  a  thin  rock  parting.  As  the 
Main  Reef  is  the  leaner,  it  has  hitherto  been  neglected  in  many  mines,  but 
it  is  expected  that  the  flushing  system  will  now  greatly  facilitate  its 
extraction  under  the  worked-out  stopes  of  the  Main  Reef  Leader. 


INDEX 


Adits,  58-59 

Advancing  system,  63,  72,  70,  79,  84,  89,  94 

retreating  system,  04,  100,  106 
Africa,  41,  63,  117 
Air,  compressed,  29-32 

drill.      See  Drill. 

fresh.      See  Ventilation. 
Altofts  Colliery,  Eng.,  67 
Amvis,  16 

Anthracite  district,  Pa.,  43,  55,  60,  110 
Appalachian  beds,  63  * 


Bacon,  Roger,  7 
Barriers.     See  Dams. 
Black  Diamond  Mine,  Pa.,  1 
Blasting,  calculations,  18 

massive  rock,  20 

seams,  26,  87 

stratified  rock,  21-24 
Bobbinite,  8 
Boyles  law,  5 
Brattices,  97-101 
Breaker,  coal,  110 
Breast-stoping,  26,  77 
Brushing.      See  Roof. 
Buggies  Mine,  81,  96 
Bulls  Head  Colliery,  Pa.,  79 
Bureau  County,  111.,  117 
Butte,  Mont.,  26,  31,  39 


Calumet  and  Hecla  Mine,  Mich.,  54 

Caps,  blasting,  2,  15 

Carbonite,  16 

Cars,  coal,  74,  85,  93,  98 

Cartridge,  3 

Charles'  law,  5 

Chocks.     See  Cribs. 

Chutes,  loading,  78,  96 

Cleaning.     See  Coal  and  Sorting. 

Cleavage,  25,  65,  84,  103 

Clod,  35 

Coal  breaker,  110 

commission,  69,  110 

cutting,   hand,   09,   74,   82,   86,    100 
machine,  26,  66,  87,  92,  100,  106,  109 

mining,  20,  35,  41,  43,  02-70,  79-117 
Cogs.     See  Cribs. 
Colliery.     See  Coal. 
Comparison  systems,  119-127 
Connellsville  district,  Pa.,  102 
Continental  Mine,  Pa.,  105 
Continuous-face.     See  Longwall. 
Conveyers,  84,  112 
Creep,  40,  75,  101 
Cribs,   75,   82,   88,   90 
Cripple  Creek,  Colo.,  28,  59 
Culm,  110,  112. 


Dams,  artificial,  55,  01 

deposition,  58 

natural,  56,   114,   115,  117 
Danville  Mines,  Pa.,  76 
Delaware  &  Hudson  Coal  Co.,  Pa.,  110 
D.  L.  &  W.  Coal  Co.,  Pa.,  110 
Development,  mine,  20 
Dip-working  of  seams,  70 
Dodson  Colliery,  Pa.,  112,   115,  117 
Dowance  Colliery,  Pa.,  114,  115,   116 
Drainage,  calculating,  47 

massive  rock,  53 


Drainage,  stratified  rock,  52,  64,  76,  90.    See  also 

Adits,  Pumps  and  Water. 
Drill,  air-hammer,  21,  24 

pointing,  20-27 

prospecting,  64,  113 
Drummond  Colliery,  N.  S.,  89 
Drywalling,  70,  75,  77,  80,  82,  114,  117 
Dust,  26,  64 
Dynamites,  13.     See  also  Blasting. 


Electric  blasting,  19,  28 

exploders,  2 

power,  86.     See  also  Haulage. 
England,  41,  67,  69,  84 
European  Mines,  03,  89,  110 
Explosions,  mine,  64,  108 
Explosives,  calculating,  4 

chemical,  8 

detonating,  2 

igniting,  1 

loading,  3 

mechanical,  7 

chemical,  12.     See  also  Blasting,  Electric, 
Powder. 


Fayal  rules,  34,  41 

Filling  seams,  75,  78,  82,  88,  100,  109,  117 

Flowage  rock,  46 

Flushing  pipe,  113 

slush,  112,  117,  118 

system,  110,  117.     See  also  Filling  and  Roof. 
Forcite,  14 
Fracture.     See  Roof. 
Frick  Coke  Co.,  Pa.,  106 
Fulminates,  11,  14 
Fuse,  2.     See  also  Blasting. 

G 

Gas,  64,  76,  91,  115 
Gelignite,  14 
Geneva,  N.  Y.,  49 
Gobbing.     See  Filling. 
Grundy  County,  111.,  72 
Gulf  of  Mexico,  49 
Guncotton,  12 


Haulage,  air,  32 

animal,  64,  74,  80,  93,  102 

electric,  98,  102 

endless  rope,  81 

tail  rope,  93 
Hausse  formulse,  42 
Hazen  formula,  51 
Hazleton,  Pa.,  110 
Headings,  driving,  20-24 
Hoisting  coal,  87,  93 
Horn  Silver  Mine,  Utah,  58 
Hudson  River  tunnels,  35 


Illinois,  72 

Inclines.     See  Jigs  and  Slopes. 

Intercoolers,  30,  32 

Iowa,  49 

Iron  Mountain  Mine,  Mont.,  55 


119 


Jeddo-Basin  adit,  Pa.,  60 
Jeffrey  Mfg.  Co.,  Ohio,  87,  112 


120 

Jigs,  haulage,  93,  90,  97 
Joplin  District,  Mo.,  26 
Joveite,  12 


Kirby,  E.  B.,  58 


Labor,  coal  mine,  63,  76,  82,  87,  99,  100 

Lake-basins,  50,  57 

Lake  Superior,  24 

La  Salle  County,  111.,  72 

Launders,  112,  117 

Lausanne  adit,  Pa.,  60 

Lehigh  Valley  Coal  Co.,  Pa.,  43,  55,  110,  116 

Leyner  drill,  21 

Locomotives.     See  Haulage. 

Longwall  system,  63,  66,  72,  76,  84,  89 

Luzerne,  Pa.,  116 

Lyddite,  12 

M 

Mahanoy  City,  Pa.,  117 

Maps.     See  Surveys. 

Merriman  formulae,  58 

Misfires,  16 

Missouri,  26 

Monongahela  River  Coal  Co.,  Pa.,  106 

Montour  Mines,  Pa.,  76 

Multi-charging,  25 

N 

Nelms,  H.  J.,  96,  100 
Newhouse  adit,  Colo.,  60 
Nitro-benzol,  15 

-gelatine,  11 

-glycerine,  9 
Nytryl,  8 
Nobel,  Alfred,  9,  11,  13 


INDEX 


o 


Oil  well.  See  Wells. 
Oneida  adit,  Pa.,  60 
Overhand  stoping,  28 


Pack-walls.     See  Drywalling. 
'-system,  longwall,  62,  79, 
pillar,  96,  100,  103 


Panel- 


110 


Pardee  Colliery,  Pa.,  110 

Pendleton  Colliery,  Eng.,  67 

Percolation.     See  Water. 

Philadelphia  &  Reading  Coal  Co.,  Pa 

Picrates,  12 

Pillar  and  stall.     See  Stall. 

-drawing,  seams,  95,   98,   104,  105,  107,  109, 
116 

-placing'  43,  72 

-system,  seams,  62,  80,  89,  94,  96,  98,  101,  102, 

106,  115 

Piping,  32,  61,  113,  116 
Pittsburgh  seam,  Pa.,  102,  106 
Plymouth  Coal  Co.,  Pa.,  110,  112 
Porter  locomotives,  32 
Portland  Mine,  Colo.,  28 
Potentite,  13 
Powder,  black,  7 

permissible,  8,  16 

smokeless,  12 
Preheater,  31 
Prop-puller,  105 
Providence,  Pa.,  79 
Pumping,  31,  50,  61,  112 

R 

Raokarock,  18 

Railroad.    See  Haulage,  Cars,  Track  and  Tunnels. 

Rainfall,  15 

Receiver,  air,  32 

Rectang,  longwall,  73 


Resistance,  line  of  least,  18 

Retreating  systems,  seams,  63,  98,  102,  106 

Richardson  formula,  42 

Rise-working,  seams,  69 

Roburite,  16 

Roof-brushing  75,  76'  90 

-control,  flat,  33-37,  65,  77 
inclined,  38,  84 

-pressure,  65-71,  75,  82,  93 

-subsidence,  40,  41,  46,  63,65,  75,  103,  106, 

107,  109,  116,  118 
Room  and  pillar.     See  Pillar. 
Roosevelt  adit,  Colo.,  59 
Run-off.     See  Water. 

S 

Scotch  longwall,  72 
Scranton,  Pa.,  WO 

Seams,  mining.  See  Advancing,  Blasting,  Coal, 
Longwall  Panel,  Pillar,  Retreating  Roo*. 
Slicing. 

overlaying,  43,  115,  118 

recovery  of,  64,  69,  94,  106,  108,  110 
Shafts,  43,  64,  72 
Shops,  72 
Siphons,  60 
Skips,  61 

Slicing  system,  117 
Slopes,  80,  89 
Slush.     See  Flushing. 
Sprengel  explosives,  15 
Spring  Valley  Collieries,  111.,  72 
Squeeze,  40,  64,  96,  98,  109 
Squibs,  1 
Stables,  95 

Stall  and  pillar,  62,  97 
Stepped-face.     See  Longwall. 
Stoping,  24-27 
Subsidence.     See  Roof. 
Sunderland,  Eng.,  41 
Support,    ground.     See    Roof,     Drywulling    and 

Timbering. 

Surface  support,  43,  115 
Surveys,  99,  101 


Tamping,  3 

Timbering,  seams,  75,  78,  87,  103 
Tonite,  3 

Top-slicing. '   See  Slicing. 
Track,  coal,  81,  93,  96,  103,  105 
Transvaal,  41,  63,  117 
Tunnels,  blast,  27 
railroad,  34,  35 

U 

Unwatering.     See  Drainage. 
V 

Ventilation,  seams,  63,  64,  80,  86,  91,  95,  96, 

100,  101,  108 
Vinton  Colliery,  Pa.,  84 


Walls.     See  Drywalling. 
Water,  diversion,  58,  58A. 

evaporation,  49 

percolation,  50 

run-off,  49 

supply,  47. 
WeiKel,  Martin,  7 
Wells,  10,  51 
Westfalit,  16 
Westville,  N.  S.,  89 
Wilkes-Barre,  Pa.,  110,  116,  117 
Williams  pulverizer,  113 
Winches,  81,  82 


Zones,  flowage  and  fracture,  46 


PRACTICAL  SHAFT  SINKING 

BY 

FRANCIS  DONALDSON,  M.  E. 


COPYRIGHT,  1910,  1912,  BY  THE 
MCGRAW-HILL  BOOK  COMPANY,  INC. 


PREFACE   TO   THE   FIRST   EDITION 

THE  subject  matter  of  this  book  was  published  as  a 
series  of  articles  in  Mines  and  Minerals,  during  1909  and 
1910.  It  is  reproduced,  with  some  alterations  and  addi- 
tions, through  the  courtesy  of  Mr.  Rufus  J.  Foster,  manager, 
and  Mr.  Eugene  B.  Wilson,  editor  of  Mines  and  Minerals. 
The  writer  also  wishes  to  acknowledge  his  indebtedness  to 
Mr.  H.  H.  Stock,  who  was  editor  of  Mines  and  Minerals 
when  most  of  the  articles  came  out. 

September,  1910. 


PREFACE   TO   SECOND   EDITION 

SINCE  the  text  of  the  first  edition  of  "Practical  Shaft 
Sinking"  was  written,  cement  grout  has  been  used  in  several 
American  shafts  to  cut  off  flows  of  water  encountered  in 
sinking,  and  its  further  use  for  this  purpose  will  undoubtedly 
become  more  common.  The  writer,  therefore,  believes  that 
a  description  of  the  methods  used  in  grouting  off  flows  of 
water  in  two  of  the  city  aqueduct  shafts  (Catskill  Aqueduct 
Project)  will  make  an  interesting  appendix.  Such  a  de- 
scription is  given  in  Appendix  A.  It  will  be  noted  that  in 
one  of  these  shafts  a  stratum  of  loose  sand  prevented  the 
entire  exclusion  of  the  water  by  grouting  alone,  that  a 
concrete  lining  provided  with  drain  pipes  was  placed,  and 
that  the  shaft  was  finally  made  entirely  dry  by  grouting 
this  lining. 

Several  of  the  city  aqueduct  shafts  in  Brooklyn  and 
the  lower  east  side  of  Manhattan  Island  were  sunk  through 
great  depths  of  water-bearing  sand  by  the  pneumatic  cais- 
son process.  A  section  of  one  of  these  caissons,  accompanied 


iv  PREFACE 

by  a  description  of  the  methods  used  in  sinking  it  and  seal- 
ing it  to  the  rock,  is  shown  in  Appendix  B. 

One  change  has  been  made  in  the  shaft  records  on  page 
83  to  accommodate  a  new  American  record.  Several  foot- 
notes have  also  been  added,  and  one  or  two  typographical 
errors  corrected. 


CONTENTS 

CHAPTER  PAGB 

PREFACE jjj 

I    SOME  DEEP  SHAFTS  —  FEATURES  OF  CONTRACTS  FOR  SINKING 

—  FORM  OF  CONTRACT 1 

II     PLANT  REQUIRED  —  BOILERS,  HOISTING  ENGINES,  HEAD-FRAME 

AND  BUCKET  —  AIR  COMPRESSORS 16 

III  SINKING  THROUGH  SURFACE  —  SOFT  GROUND  —  WOODEN  SHEET- 

ING —  STEEL   SHEETING  —  CAISSONS   OF   STEEL,    WOOD,    OR 
CONCRETE     •,  .  .  ' 33 

IV  SINKING    THROUGH    SOFT    GROUND  —  PNEUMATIC     PROCESS  — 

SHIELD  METHOD 59 

V    SINKING   IN   ROCK  —  ARRANGEMENT   OF   HOLES  —  TOOLS   AND 

METHODS  USED  IN  DRILLING  —  COSTS  AND  SPEED      ...       66 
VI    THE  SINKING-DRUM  PROCESS  —  MAMMOTH  PUMP  —  THE  FREEZ- 
ING PROCESS 85 

VII    THE    KIND-CHAUDRON    BORING    PROCESS  —  CEMENTATION    OF 

WATER-BEARING  FISSURES .     .     .     .       97 

VIII     LIFTING  WATER  —  HORIZONTAL  vs.  VERTICAL  PUMPS  —  HAND- 
LING PUMPS  IN  SHAFT  —  CORNISH  PUMPS    .      .      .      .      .      .     107 

IX     SHAFT  LININGS '.....     118 

X    CONCRETE  LININGS  —  COSTS  PER  LINEAR  FOOT  FOR  RECTANGU- 
LAR ELLIPTICAn,  AND  QUADRILATERAL  SHAFTS 129 

APPENDIX  A.     GROUTING  SHAFTS  4  AND  24,  N.Y.  CITY  AQUEDUCT     140 
B.  .....' 144 


PRACTICAL  SHAFT  SINKING 

CHAPTER  I 

SOME  DEEP  SHAFTS  —  FEATURES  OF  CONTRACTS  FOR 
SINKING — FORM  OF  CONTRACT 

THE  origin  of  mining  is  lost  in  the  mists  of  antiquity, 
but  it  is  certain  that,  since  the  beginning  of  history,  metals 
and  minerals  have  been  sought  after.  The  Egyptians 
operated  gold,  silver,  and  copper  mines  in  Ethiopia  and  on 
the  Arabian  border;  the  Phoenicians  found  gold  and  iron 
in  the  islands  of  the  Mediterranean  and  lead  and  silver 
in  Spain.  The  earliest  mines  were  probably  surface  work- 
ings, but  the  first  historical  mention  of  openings  driven  in 
the  earth  refers  not  to  a  drift  or  tunnel  but  to  a  shaft.  In 
the  Book  of  Job  it  is  written  of  man  that  "He  breaketh 
open  a  shaft  away  from  where  men  sojourn;  they  are  for- 
gotten of  the  foot;  they  hang  afar  from  men;  they  swing  to 
and  fro."  Pliny  describes  cutting  hitches  in  a  shaft:  "Else- 
where pathless  rocks  are  cut  away  and  are  hollowed  out  to 
furnish  a  rest  for  beams.  He  who  cuts  is  suspended  with 
ropes." 

Shaft  sinking  and  tunnel  operations  in  ancient  times 
were  confined  to  solid  earth  and  rock.  The  Roman  engi- 
neers drove  rock  tunnels  that  would  seem  long  to-day;  they 
originated  the  method  of  disintegrating  rock  by  fire  and  they 
sunk  shafts  along  the  line  of  their  tunnels  from  which  to 
drive  additional  headings.  Forty  shafts  —  one  of  them 
400  ft.  deep  —  were  used  for  the  excavation  of  their  longest 
tunnel. 

For  many  centuries  after  the  Roman  Era  nothing  com- 
parable to  the  Roman  work  was  attempted,  since  the  cost 
in  labor  and  human  life  of  the  fire-and-water  method  was 

i 


2  PRACTICAL  SHAFT  SINKING 

terrific.  The  invention  of  gunpowder  was  the  next  step,  but 
gunpowder  was  apparently  not  used  for  blasting  purposes 
until  1679,  at  Malpas,  France.  Mines  in  the  Hartz  Moun- 
tains and  in  Cornwall  had  been  worked  to  great  depths 
in  the  seventeenth  century  before  the  steam  engine  was 
developed,  but  its  application  to  hoisting  of  course  made 
possible  undreamed  of  speed  in  sinking.  The  first  practical 
use  of  steam  was,  incidentally,  to  pump  water  from  the  Cor- 
nish shafts. 

The  invention  of  dynamite,  the  first  commercial  high 
explosive,  in  1866,  and  the  compressed-air  drill  in  1855,  put 
rock  shaft  sinking  on  its  present  basis.  Although  from 
time  to  time  special  methods  such  as  the  freezing  and  the 
boring  processes  have  been  developed  for  special  conditions, 
for  ordinary  shafts  hand  sinking  is  cheapest  and  best. 
Excepting  the  steam  hoist,  inventions  have  been  confined 
to  means  for  shattering  the  rock;  steam  shovels  are  some- 
times used  in  tunnels,  but  shaft  spoil  is  to-day  loaded  by 
hand  into  buckets,  as  in  the  days  of  the  Romans. 

Before  the  last  half  of  the  nineteenth  century,  soft- 
ground  sinking  was  confined  to  material  penetrable  by  fore- 
poling.  Although  considerable  depths  have  been  reached 
in  this  way,  where  the  ground  is  bad  the  method  is  at  best 
slow  and  precarious.  The  Germans  originated  the  hydrau- 
lically  forced  sinking  drum  and  the  freezing  process.  The 
pneumatic  process  was  first  used  by  Brunei  in  the  Thames 
Tunnel.  Recently,  concrete  sinking  drums  or  open  caissons 
have  been  extensively  used. 

The  sizes  and  shapes  of  shafts  are  governed  by  the  nature 
of  the  material  to  be  hoisted  through  them,  by  the  char- 
acter of  the  ground  to  be  penetrated,  and  also  largely  by  local 
usage.  Since  mine  cars  and  skips  are  approximately  rect- 
angular in  plan,  a  rectangle  is  the  most  economical  shape 
for  a  hoist  shaft,  giving  the  maximum  usable  area  with  the 
minimum  excavation;  this  advantage,  however,  does  not 
apply  to  an  air  shaft.  The  rectangular  shape  is  also  adapted 
to  timbering,  the  cheapest  form  of  lining,  and  is  on  this 


SOME  DEEP  SHAFTS  3 

account  standard  in  America.  In  Europe,  on  the .  other 
hand,  all  shafts  are  circular  or  elliptical  and  are  lined  with 
brick  or  concrete  masonry.  This  type  has  the  disadvantage 
of  high  first  cost,  but  a  masonry  lining  is  proof  against 
decay  and  fire  and  explosions.  In  wet  strata  also,  a  circular 
shaft  may  be  lined  with  iron  tubbing  and  thus  kept  entirely 
dry. 

In  large  mines  two  openings  are  always  advisable  to 
secure  satisfactory  ventilation;  in  coal  mines  where  explo- 
sive gases  form  they  are  absolutely  necessary,  and  in  most 
states  are  required  by  law.  The  hoist  shaft  may  be  upcast 
or  downcast;  in  either  case  an  airway  is  usually  provided 
in  addition  to  the  hoist  compartments.  All  mines  worthy 
of  the  name  have  balanced  cages  requiring  two  hoistways; 
the  airway  makes  a  three-compartment  shaft  the  most 
common  type.  In  rectangular  shafts,  where  several  com- 
partments are  needed,  a  long  shaft  one  compartment  wide 
is  easier  to  sink  and  timber  than  a  short  shaft  two  com- 
partments wide;  for  instance,  if  four  7  X  10  ft.  compart- 
ments are  desired,  a  shaft  10  X  28  ft.  is  preferable  to  one 
20  X  14  ft. 

In  America,  in  the  bituminous  coal  fields,  hoist  shafts 
are  usually  13  X  26  ft.  in  the  rock,  are  lined  with  8  X  10  in. 
timber  and  have  two  7  X  11  ft.  hoistways  and  a  9  X  11  ft. 
airway.  In  wet  mines  a  5-ft.  pipeway  is  added.  The  air 
shafts  are  13  X  18  ft.,  with  a  10  X  11  ft.  airway  and  a  6-ft. 
stairway  compartment.  In  the  anthracite  fields  the  deeper 
hoist  shafts  sometimes  have  four  hoistways  operating  from 
several  coal  seams,  besides  air  and  pipe  ways,  and  have 
sections  12  X  42  ft.  to  14  X  56  ft.  in  the  rock.  European 
coal  shafts  are  customarily  20  to  23  ft.  in  finished  diameter. 
Coal  shafts  are  almost  always  vertical. 

In  ore  mines  different  conditions  prevail.  Ore  is  less 
bulky  than  coal,  is  harder  to  mine,  and  can  be  loaded  through 
chutes  without  objectionable  breakage.  Large  shafts  are 
therefore  unnecessary  and  the  sizes  range  from  7  X  9  ft. 
in  the  iron  mines  formerly  operated  at  Boyertown,  Pa., 


4  PRACTICAL  SHAFT   SINKING 

to  .9  X  24  ft.  in  the  Michigan  iron  country.  Ore 
shafts  are  usually  sunk  on  the  vein,  and  so  may  be  found 
at  any  inclination  with  the  vertical,  but  where  natural 
conditions  do  not  compel  an  inclination,  a  perpendicular 
shaft  is  preferable. 

The  deepest  shaft  in  America,  No.  3  Tamarack  at  Tama- 
rack, Mich.,  is  5253  ft.  deep  and  is  used  in  mining  copper. 
No.  5  shaft  at  Tamarack  is  5180  ft.  deep,  Red  Jacket  shaft 
at  Calumet,  Mich.,  is  4900  ft.  deep.  These  shafts  are 
remarkable  not  only  because  they  penetrate  the  earth  for 
almost  a  mile,  but  also  because  of  the  remarkably  powerful 
hoisting  engines  used  —  engines  which  hoist  a  total  load  of 
17  tons  at  the  rate  of  6000  ft.  per  minute.  All  of  these  shafts 
are  vertical. 

In  the  Pennsylvania  anthracite  fields,  where  acid  mine 
water  quickly  eats  up  pumps  and  piping,  a  number  of  shafts 
have  been  sunk  for  the  purpose  of  hoisting  water.  The 
tanks  used  for  hoisting  fill  and  empty  themselves  auto- 
matically, discharging  the  water  into  a  basin  at  the  top 
of  the  shaft.  Powerful  hoist  engines  are  provided.  The 
most  notable  shaft  of  this  type  is  owned  by  the  D.  L. 
&  W.  R.  R.,  at  Scranton,  Pa.  It  is  entirely  automatic, 
requiring  no  engineer,  and  is  operated  through  friction 
clutches  by  an  800-horse-power  induction  motor. 

The  driving  of  rock  shafts  and  tunnels  is  very  unlike  the 
mining  of  coal;  a  different  class  of  workmen,  different  fore- 
men, and  different  tools  are  needed.  It  is  seldom  that  a 
good  coal-mine  foreman  is  also  a  good  sinker,  and  good 
sinkers,  unattached,  are  not  always  easy  to  obtain.  For 
these  reasons  it  is  customary  for  coal-mining  companies  to 
have  a  large  part  of  their  development  work  done  by  con- 
tract, and  even  the  large  anthracite  corporations,  who  own 
the  necessary  surface  equipment  for  sinking,  prefer  to  have 
contractors  do  the  sinking.  In  ore  mines  the  foregoing  does 
not  apply;  sinking  shafts  is  part  of  the  day's  work  and  all 
the  miners  are  rock  men,  but  in  opening  a  new  mine  the 
question  of  time  is  still  to  be  considered.  The  loss  of 


DISPOSAL    OF    SOIL  5 

interest  on  the  investment  in  a  large  property  before  it  is 
developed  may  amount  to  several  hundred  dollars  a  day, 
and  every  day  lost  in  sinking  adds  that  much  to  the  cost 
of  the  shaft.  Even  when  a  mining  company  is  so  situated 
that  it  can  sink  its  shaft  cheaply,  a  responsible  contractor 
possessing  a  plant  and  an  organization  can  save  enough 
time  to  more  than  pay  his  profit. 

When  it  is  decided  to  have  a  shaft  sunk  by  contract, 
the  first  essential  is  to  get  trustworthy  contractors  to  bid 
on  the  job;  the  second  is  to  prepare  a  contract  fair  to  both 
sides.  While  it  is  not  well  to  leave  loopholes  whereby  the 
contractor  can  escape  from  the  plain  provisions  of  the  speci- 
fications, it  is  equally  unwise  to  attempt  to  tie  him  down  so 
tight  in  every  detail  that  he  is  practically  dared  to  find  a 
flaw  in  the  agreement.  It  is  almost  impossible  to  foresee 
every  contingency,  and  an  omission  in  a  very  tightly-drawn 
contract  is  harder  to  correct  subsequently  than  an  omission 
in  a  looser  one. 

A  complete  shaft  contract  form  may  be  found  in  several 
text-books,  or  obtained  elsewhere  without  difficulty;  a  form 
in  common  use  is  appended  to  this  chapter.  Among  the 
specific  points  that  warrant  attention  may  be  mentioned 
the  following: 

Disposal  of  Spoil.  —  The  labor  cost  of  a  shaft  is  of 
course  directly  affected  by  the  nature  of  the  dump.  It  is 
also  indirectly  affected  by  it  to  an  even  greater  extent. 
The  delays  to  sinking  caused  by  a  long  haul  and  an  incon- 
venient dump,  especially  in  bad  weather,  are  likely  to  be 
more  serious  than  the  cost  of  the  additional  labor  required. 
The  specifications  should,  therefore,  state  where  the  spoil 
is  to  be  dumped,  or  at  least  where  it  is  not  to  be  dumped. 
In  one  case  where  a  contract  contained  the  usual  clause,  "the 
spoil  shall  be  placed  where  the  engineer  shall  direct,  the  haul 
not  to  exceed  1500  ft.,"  no  plans  were  available,  and  in  the 
absence  of  any  direct  prohibition  by  the  engineer,  the  con- 
tractor started  to  dump  spoil  in  the  vicinity  of  the  shaft. 
After  three  months'  sinking,  the  engineer  discovered  that 


6  PRACTICAL  SHAFT  SINKING 

the  dump  was  in  his  way  and  directed  that  the  spoil  be 
placed  elsewhere.  Subsequently  he  compelled  the  contrac- 
tor, under  the  clause  cited  above,  to  move  all  the  spoil 
dumped  in  the  first  three  months.  While  in  this  case  the 
engineer's  order  might  have  been  successfully  contested, 
much  trouble  could  have  been  saved  by  proper  care  in 
drawing  up  the  contract. 

Time  Limit  and  Penalty.  —  A  time-limit  clause  is  most 
properly  a  feature  of  nearly  all  sinking  contracts,  and  the 
usual  provision  made  to  secure  its  enforcement  is,  "and  in 
the  event  of  the  contractor  failing  to  complete  the  work  by 
this  date,  it  is  mutually  agreed  that  he  shall  pay  the  con- 
tractee  the  sum  of  -  -  dollars  for  every  day  thereafter 
until  the  work  is  completed,  not  as  a  penalty,  but  as  liqui- 
dated damages."  In  spite  of  this  definition,  the  courts  have 
often  held  that  the  actual  damages  must  be  proven  and  the 
possibility  of  collecting  the  stated  damages  is  not  assured. 
Since  the  real  damages  to  a  mining  company  due  to  delay 
in  getting  started  consist  of  the  loss  of  interest  on  the  invest- 
ment, in  every  case  where  the  company  has  its  surface 
arrangements  ready  for  work  before  its  shafts  are  finished, 
it  will  gain  as  much  per  day  by  their  completion  ahead  of 
time  as  it  will  lose  by  their  non-completion.  The  writer 
therefore  believes  that  where  a  penalty  is  to  be  collected  for 
delay  an  equal  premium  should  be  paid  for  time  saved,  not 
only  because  this  is  fair  but  also  because  it  is  likely  to 
expedite  the  work.  An  extension  of  time  is  usually,  and 
should  be,  allowed  the  contractor  on  account  of  "  unusual 
difficulties  with  water  or  quicksand." 

Acceptance.  —  Where  two  shafts  are  sunk  simultaneously 
under  the  same  contract  (as  in  opening  a  new  coal  mine), 
the  first  shaft  down  is  usually  accepted  by  the  company  upon 
completion.  If  this  is  not  the  intention  it  should  be  so 
stated  in  the  contract. 

Risk  of  Water.  —  It  has  been  the  practice  in  shaft  con- 
tracts to  throw  the  risk  of  encountering  unusual  quantities 
of  water  upon  the  contractor.  With  a  fixed  price  per  foot 


SUPPLIES  AND  MACHINERY  7 

for  sinking,  based  upon  usual  conditions,  the  contractor  will 
lose  money  if  he  strikes  water  exceeding,  say,  150  gallons  per 
minute.  With  greater  quantities  his  loss  is  often  measured 
only  by  his  financial  resources.  This  puts  a  responsible 
contractor  at  a  disadvantage,  especially  in  a  new  territory, 
for  since  he  has  the  equipment  and  the  money  needed  to 
fight  large  quantities  of  water,  he  must  raise  his  price  on  all 
shafts  to  insure  him  against  an  occasional  heavy  loss.  An 
irresponsible  contractor,  on  the  other  hand,  having  little 
to  lose,  can  afford  to  bid  low,  and  if  he  does  strike  much 
water  abandon  the  job.  The  water  risk  belongs  properly 
not  to  the  contractor  nor  the  mine  owners  in  general,  but  to 
the  owners  of  the  particular  mine  in  question ;  for  this  reason 
a  water  clause  is  a  feature  of  many  recent  contracts.  The 
New  York  Board  of  Water  Supply,  in  its. contracts  for  the 
construction  of  the  inverted  siphons  on  the  Catskill  aque- 
duct, calls  for  a  price  for  pumping  each  million  gallons  of 
water  1  ft.  In  these  jobs  the  time  schedules  are  so  carefully 
worked  out  that  there  is  little  likelihood  of  the  contractor 
pumping  water  for  profit,  but  a  form  of  water  clause  more 
acceptable  to  the  average  mining  company  is  one  in  which 
the  contractor  makes  an  additional  price  per  foot  for  every 
hundred  gallons  per  minute  pumped  while  sinking.  He  is 
thus  paid  nothing  if  the  shaft  is  idle  and  is  encouraged  to 
make  progress. 

Supplies  and  Machinery.  —  Local  conditions  determine 
whether  the  mining  company  or  the  contractor  should 
furnish  the  supplies  or  machinery.  The  larger  anthracite 
companies,  who  hold  extensive  properties  and  open  new 
mines  upon  them  as  the  need  arises,  own  sinking  engines, 
boilers,  and  other  equipment;  at  a  new  shaft  they  erect  a 
surface  plant  complete,  furnish  timber  and  coal,  and  expect 
the  contractor  to  supply  only  the  air  compressor  and  drills, 
small  tools,  and  labor.  In  the  bituminous  fields  and  in 
many  ore  regions,  the  mining  companies  usually  wish  to 
develop  new  property  as  soon  as  it  is  acquired,  and,  in  order 
to  concentrate  their  efforts  upon  the  permanent  surface 


8  PRACTICAL  SHAFT  SINKING 

plant,  expect  the  shaft  contractor  to  furnish  everything 
he  needs.  A  quicker  start  can  be  made  in  this  way.  If  the 
shaft  is  to  be  timbered,  however,  the  timber  should  be 
furnished  by  the  company.  Oak  suitable  for  shaft  linings 
is  rapidly  becoming  unobtainable;  yellow  pine  must  be 
brought  long  distances  and  is  not  likely  to  arrive  too  soon 
if  ordered  when  the  shaft  contract  is  let.  Aside  from  the 
question  of  speed,  the  company  by  supplying  timber  will 
save  itself  the  profit  that  the  contractor  adds  to  cost,  and 
also  occasional  vexatious  squabbles  as  to  quality. 

No  matter  how  the  shaft  is  sunk  the  permanent  boiler 
plant  should  be  made  ready  to  operate  as  soon  as  is  prac- 
ticable. The  possibility  of  striking  water  is  the  greatest 
hazard  of  sinking  and,  if  water  is  encountered,  the  first 
requirement  is*  plenty  of  steam.  It  is  easier  and  quicker  to 
procure  and  install  any  amount  of  pipe  and  any  number 
of  pumps  than  it  is  to  build  a  boiler  plant  to  run  the  pumps. 
Effective  sinking  pumps  are  so  exceedingly  wasteful  of 
steam  that  the  permanent  mine  boiler  plant  will  be  none 
too  large  and  efficient  to  care  for  a  large  inflow  of  water;  its 
early  completion  will  insure  against  a  long  and  costly  delay. 

The  local  flow  of  underground  water  is  one  of  the  most 
uncertain  features  of  geology,  and  whether  or  not  water 
will  be  encountered  in  a  given  shaft  can  seldom  be  predicted 
with  certainty.  Even  when  borings  at  the  site  of  a  shaft 
or  other  shafts  sunk  in  the  vicinity  indicate  that  water  will 
be  struck,  the  amount  is  problematical.  A  few  general 
remarks  may  be  stated  as  follows: 

Since  the  rainfall  in  mountains  is  high,  the  ground  near 
them  will  have  an  opportunity  to  collect  water.  Geologic 
faults  form  passages  whereby  surface  water  finds  its  wray 
into  the  ground,  and  the  fissures  caused  by  the  strain  to 
which  the  rock  was  subjected  when  the  fault  was  made  act 
as  reservoirs.  A  shaft  on  or  near  a  fault  is  sure  to  be  wet 
below  the  ground-water  level  of  the  surrounding  region. 
Natural  water  courses  also  form  in  soluble  rock  such  as 
sandstone  and  limestone,  especially  the  latter.  An  example 


SOME  DEEP  SHAFTS  9 

of  a  water  course  of  this  kind  was  afforded  at  the  zinc  mine 
at  Friedensville,  Pa.,  formerly  drained  by  the  "  President," 
the  largest  Cornish  pump  ever  built.  This  mine  is  located 
at  the  foot  of  a  mountain. 

Two  shafts  sunk  in  the' Allegheny  Mountains  near  South 
Fork,  Pa.,  also  encountered  a  water  course.  They  were 
about  250  ft.  apart  and  apparently  were  sunk  directly  on 
top  of  the  water  channel.  What  may  be  called  the  down- 
stream shaft  was  sunk  first  through  the  wet  stratum;  as  the 
up-stream  shaft  was  sunk  the  flow  into  it  increased,  while  the 
flow  into  the  down-stream  shaft  decreased  at  the  same  rate. 

Shafts  near  streams  are  likely  to  strike  water  at  the  sur- 
face of  the  rock,  but  not  necessarily  below  it  if  the  rock  is 
solid.  In  ordinary  coal  measures  a  feeder  may  be  expected 
at  the  seam  between  an  upper  permeable  rock  like  sandstone, 
and  a  lower  bed  of  impervious  shale  or  fireclay. 

The  uses  of  the  diamond  drill  in  prospecting  for  coal  and 
ore  are  too  well  known  to  require  comment,  as  far  as  the 
knowledge  obtained  of  the  rock  is  concerned.  A  hole  near 
a  proposed  shaft  will  also  give  much  information  as  to  the 
ground-water  conditions,  even  though,  as  has  been  said, 
the  quantity  cannot  be  determined.  A  diamond  drill  hole 
is  not  large  enough  to  pump  out,  but  the  process  may 
be  reversed.  If  additional  water  can  be  pumped  into  a 
hole  already  full,  the  strata  are  evidently  open  enough  to 
let  water  into  a  shaft.  A  bore  hole,  of  course,  may  be 
pumped  with  a  deep-well  pump  or  air  lift;  it  has  in  fact 
been  suggested  that  wet  ground  be  drained  by  pumping 
from  a  ring  of  bore  holes  around  the  shaft  location,  thus 
doing  away  entirely  with  pumps  in  the  shaft. 

Prospect  holes  should  be  located  at  one  side  of  the  shaft, 
so  that  if  a  pocket  of  water  is  drilled  into  at  a  considerable 
depth,  it  will  not  rise  into  the  shaft  through  the  hole.  In 
this  way  pumps  and  piping  need  not  be  installed  until  the 
bottom  of  the  shaft  has  almost  reached  the  level  of  the 
pocket,  and  the  depth  of  the  wet  sinking  is  reduced  to  a 
minimum. 


10  PRACTICAL  SHAFT  SINKING 


CONTRACT   AGREEMENT   FOR   SHAFT   SINKING 

This  agreement  made  in  duplicate  this         day  of 
by  and  between  the  Coal  Co.,  a  corporation  chartered 

and  existing  under  the  laws  of  the  State  of 
party  of  the  first  part,  and  the  Con- 

tracting Co.,  a  corporation  chartered  and  existing  under  the 
laws  of  the  State  of  ,  party  of  the  second  part, 

WITNESSETH,  That  for  and  in  consideration  of  the  cove- 
nants and  payments  hereinafter  specified  to  be  made  and  per- 
formed by  the  party  of  the  first  part,  the  said  party  of  the 
second  part  doth  hereby  covenant  and  agree  to  build  and 
complete  in  the  most  substantial  and  workmanlike  manner,  a 
hoisting  shaft  13  X  26  ft.,  outside  the  timbers,  and  an  air 
shaft  13  X 18  ft.,  outside  the  timbers,  each  approximately  600 
ft.  deep,  for  the  Coal  Co.,  at  its  property  near 

.  The  work  is  to  be  done  in  accordance  with  the 
attached  specifications  and  the  plans  furnished  by  the  party 
of  the  first  part,  which  are  hereby  made  a  part  of  this  agree- 
ment; the  party  of  the  second  part  is  to  furnish  all  the  labor 
and  materials  necessary,  except  such  as  are  particularly 
noted  in  the  specifications  as  being  furnished  by  party  of  the 
first  part. 

The  said  work  is  to  be  completed  on  or  before  the  first 
day  of  ,  19  . 

And  the  party  of  the  first  part  doth  promise  and  agree 
to  pay  the  party  of  the  second  part  the  following  prices  for 
the  several  kinds  of  work  herein  specified,  of  which  the 
'following  is  a  summary: 

HOIST  SHAFT 

Excavation  measured  from  top  of  natural  ground  to  bottom  of 

coal  seam $  per  vert.  ft. 

Framing  and  placing  all  timber  and  lagging $  per  vert.  ft. 

Water  rings  complete    $  each. 

AIR  SHAFT 
Same  as  above. 


CONTRACT  AND  SPECIFICATIONS  11 

The  above  prices  contemplate  a  maximum  of  not  more 
than  100  gallons  of  water  per  minute  to  be  pumped  from 
each  shaft. 

The  following  extra  prices  will  be  paid  in  each  shaft 
for  each  100  gallons  per  minute  in  excess  of  this  amount,  as 
follows : 

Water  Pumped;  Additional  Price 

Gallons  per  Min.  Paid  per  Foot 

100  -  200 $  15 

200  -  300 30 

300  -  400 45 

400  -  500 60 

500  -  600 75 

600  -  700 90 

700  -  800 110 

800  -  900 130 

900  -  1000 150 

If  the  volume  of  water  should  exceed  1,000  gallons  per 
minute,  a  supplementary  agreement  will  be  made. 

The  payment  for  said  work  shall  be  made  in  the  following 
manner : 

An  estimate  will  be  made  about  the  last  day  of  every 
month  of  the  amount  of  work  done  during  the  month,  and 
90  per  cent,  of  the  same  will  be  paid  on  or  before  the  20th 
of  the  succeeding  month,  10  per  cent,  of  the  total  amount 
being  retained  until  the  entire  completion  of  the  work. 

And  when  all  the  work  embraced  in  this  contract  is 
completed,  the  party  of  the  first  part  shall,  upon  notification 
from  the  party  of  the  second  part,  make  a  final  inspection; 
if  the  work  is  found  to  be  in  accordance  with  the  specifica- 
tions, there  shall  be  a  final  estimate  made  of  the  value  of 
said  work,  according  to  the  terms  of  this  agreement.  The 
balance  due  the  party  of  the  second  part  shall  be  paid 
within  thirty  days  thereafter,  upon  said  contractor  giving 
a  release  under  seal  to  the  party  of  the  first  part  from  all 
claims  or  demands  whatsoever  growing  in  any  manner  out 
of  this  agreement;  and  upon  said  contractor  delivering  to 
party  of  the  first  part  full  release  in  proper  form  and  duly 
executed  of  all  liens,  claims,  or  demands  from  mechanics 


12  PRACTICAL  SHAFT  SINKING 

and  material  men  for  work  done  on  or  about  the  shafts,  or 
for  materials  furnished  for  the  work  under  this  contract. 

It  is  further  agreed  between  said  parties  that  said  party 
of  the  second  part  shall  not  transfer  or  sublet  any  part  of 
this  contract  to  any  person  (except  for  delivery  of  materials) 
without  the  consent  of  the  party  of  the  first  part,  and  that 
the  party  of  the  second  part  will  at  all  times  give  personal 
attention  to  the  superintendence  of  the  work. 

It  is  further  agreed  that  the  work  embraced  in  this 
contract  shall  be  commenced  within  two  (2)  weeks  of  the 
date  of  this  contract  and  prosecuted  day  and  night  (except 
Sundays)  with  as  many  men  as  can  be  worked  to  advantage. 
If,  during  the  progress  of  the  work,  it  is  the  opinion  of  the 
Engineer  of  the  party  of  the  first  part  that  the  party  of  the 
second  part  is  not  furnishing  materials  or  appliances  or  labor 
of  the  right  quality,  or  in  sufficient  quantity  to  complete  the 
work  within  the  time  agreed  on,  the  said  Engineer  may  in 
either  or  both  of  the  above-mentioned  cases  purchase  such 
material  and  appliances  or  employ  such  labor  as  in  his 
judgment  may  be  necessary.  And  the  said  Engineer  is 
authorized  to  pay  such  wages  for  labor  and  such  prices  for 
materials  and  appliances  as  may  be  found  necessary  or 
expedient,  and  to  deduct  the  amount  so  paid  from  any 
moneys  due  the  party  of  the  second  part  from  the  party  of 
the  first  part. 

It  is  further  agreed  that  the  Engineer  shall  have  the 
authority  to  order  any  additional  work  or  materials  not 
called  for  in  the  plans  and  specifications  that  he  may  deem 
necessary  or  advisable,  but  in  case  any  such  extra  work  or 
materials  is  required,  the  same  shall  be  ordered  by  the 
Engineer  in  writing,  and  the  price  for  said  extra  work  or 
materials  shall  mutually  be  agreed  upon  in  writing  before 
said  materials  are  furnished  or  said  work  is  done. 

IN  WITNESS  WHEREOF  the  parties  herein  have  hereunto 
set  their  hands  and  seals,  the  day  and  date  first  above 
mentioned. 


CONTRACT  AND  SPECIFICATIONS  13 

SPECIFICATIONS  FOR  SINKING  AND  LINING  Two  SHAFTS 
FOR  THE  COAL  COMPANY,  AT 

GENERAL 

Meaning  of  Titles.  —  The  word  Contractor,  when  here- 
inafter used,  shall  refer  to  the  Contracting  Company 
as  in  the  attached  Agreement.  The  word  Engineer  shall 
refer  to  the  Chief  Engineer  of  the  Coal  Company, 
or  his  representative. 

Labor  and  Materials  Furnished. —  The  Contractor  shall 
furnish  all  machinery,  tools,  labor,  materials,  and  supplies 
incidental  to,  or  in  any  way  connected  with,  the  sinking  and 
timbering  of  the  two  shafts  hereinafter  described,  with  the 
exception  of  the  timber  which  will  be  furnished  by  the 
Coal  Company  free  on  board  cars  at 

Location  of  Temporary  Plant.  —  The  Contractor's  hoist- 
ing apparatus  and  temporary  machinery  and  buildings  shall 
be  so  placed  as  not  to  interfere  with  the  construction  of  the 
permanent  head-frames,  or  the  erection  of  the  permanent 
plant. 

EXCAVATION 

The  dimensions  of  the  hoist  shaft  shall  be  13  X  26  ft. 
and  of  the  air  shaft  13  X  18  ft.,  outside  of  lagging.  The 
excavation  shall  be  carried  down  square  and  plumb  from  top 
to  bottom  and  be  large  enough  to  give  room  for  the  proper 
wedging  of  the  timber.  Special  care  must  be  exercised  in 
blasting  to  avoid  shattering  the  walls  of  the  shaft,  and  all 
loose  material  which  might  endanger  the  timbering  or  the 
men  working  below  must  be  removed. 

The  Contractor  shall  keep  his  machinery  and  tools  in 
good  condition,  and  take  every  reasonable  precaution  to 
insure  the  safety  of  his  men. 

The  Contractor  shall  deposit  all  material  excavated 
from  the  shafts  at  places  directed  by  the  Engineer,  to  con- 
form to  the  grades  established  adjacent  to  the  shaft.  Any 
overhaul  exceeding  500  ft.  shall  be  paid  for  at  the  rate  of 
cents  per  cubic  yard  for  each  100  ft.  of  overhaul. 


14  PRACTICAL  SHAFT  SINKING 

The  depth  of  the  shafts  shall  be  measured  from  the 
elevation  of  the  original  surface  of  the  ground  in  the  center 
of  the  shaft  to  the  bottom  of  the  coal  seam. 

TIMBERING 

The  shaft  shall  be  timbered  throughout  with  sound  oak 
or  yellow  pine  to  be  furnished  by  the  Coal  Company. 

It  is  to  be  framed  accurately  by  the  Contractor  according 
to  the  Coal  Company's  plans  and  shall  be  placed  in  the 
shafts  square,  level,  and  to  line. 

Wall  plates,  end  plates,  buntons,  and  posts  shall  be 
8  X  10  in.,  and  bearing  or  hitch  timbers  shall  be  8  X  12  in. ; 
lagging  shall  be  of  2-in.  plank.  The  lagging  shall  rest  on  a 
2  X  4  in.  oak  piece  placed  horizontally  in  the  middle  of  the 
back  of  each  end  and  wall  plate  and  well  spiked ;  the  space 
between  the  lagging  and  the  rock  shall  be  backed  solid  with 
sound  slabs  or  other  sound  refuse  timber. 

Each  corner  of  each  ring  of  timbers  and  each  wall  plate 
at  both  ends  of  every  bunton  shall  be  thoroughly  braced 
against  the  side  of  the  shaft  by  blocks  and  wedges.  The  sets 
of  timber  shall  be  5  ft.  apart  vertically,  center  to  center. 
At  intervals  of  50  ft.  vertically,  bearing  or  hitch  timbers 
shall  be  placed  to  serve  as  supports  to  the  timbering  above. 
The  hitch  or  bearing  in  the  rock  at  each  end  of  each  timber 
shall  be  strong  enough  to  develop  the  full  strength  of  the 
timber;-  in  no  case  shall  it  be  less  than  8  in.  in  depth.  The 
intervals  of  50  ft.  may  in  the  judgment  of  the  Engineer  be 
varied,  but  no  such  variations  shall  be  made  by  the  Con- 
tractor without  the  consent  of  the  Engineer. 

The  timbering  shall  be  carried  above  the  natural  ground 
to  the  level  indicated  by  the  Engineer.  Special  timbering 
shall  be  placed  at  the  shaft  bottom  in  accordance  with  the 
plans  furnished. 

The  air  compartment  shall  be  lined  with  1-in.  tongued 
and  grooved  yellow  pine  flooring,  free  from  knots  and  well 
matched  and  joined  on  end  and  wall  plates.  The  guides 
shall  be  6  X  8  in.  yellow  pine  surfaced  on  all  faces,  and  shall 


CONTRACT  AND  SPECIFICATIONS  15 

be  framed  as  shown  on  plan  and  placed  exactly  plumb,  and 
straight  and  true  to  gage  from  top  to  bottom. 

WATER   RINGS 

The  water  rings  shall  be  constructed  before  the  timbering 
is  finally  placed,  and  shall  be  as  shown  on  the  plans.  The 
bottom  of  each  ring  shall  have  a  water-tight  floor  of  concrete. 
The  number  and  location  of  the  water  rings  shall  be  deter- 
mined by  the  Engineer. 

USE  OF  CONTRACTOR'S  PLANT 

The  Coal  Company  shall  have  the  privilege  of  renting 
the  Contractor's  hoisting  and  pumping  plant  after  the  com- 
pletion of  the  shafts  for  a  period  of  two  (2)  weeks.  It  shall 
pay  the  Contractor  $  per  day  as  rental  for  said  plant. 


CHAPTER  II 

PLANT  REQUIRED  —  BOILERS,  HOISTING  ENGINES,  HEAD- 
FRAME  AND  BUCKETS  —  AIR  COMPRESSORS 

PLANT 

IN  considering  the  subject  of  shaft  sinking  from  the 
mechanical  side,  the  first  and  most  important  consideration 
is  the  proper  design  and  arrangement  of  the  surface  plant. 
The  underground  plant  comprises  rock  drills  and  pumps, 
and  both  above  and  below  ground  many  tools  and  con- 
trivances are  required.  The  items  included  under  surface 
plant  will  be  treated  first  and  the  underground  contrivances 
taken  up  later  in  connection  with  the  work  which  they 
perform. 

The  elements  of  a  modern  surface  plant  are:  Primary- 
power  producer;  hoisting  apparatus;  secondary-power  pro- 
ducers; buildings,  shops,  etc. 

Primary-power  Producer.  —  Although  in  a  few  favored 
localities  electric  power  may  be  cheaply  bought  and  used 
directly,  or  converted  into  air  power  when  needed,  in  nine 
tenths  of  the  shafts  sunk  the  primary  power  is  steam.  The 
boiler  plant,  in  this  case,  is  the  backbone  of  the  job;  it  must 
be  put  up  to  allow  of  expansion  if  necessary,  and  it  must  be 
absolutely  reliable.  In  other  forms  of  construction  work, 
water,  while  always  a  source  of  trouble  and  expense,  is  not 
the  implacable  enemy  that  it  is  in  sinking.  The  pumps 
which  drain  a  cofferdam  will  also  serve  to  empty  it,  and  a 
breakdown  delays  the  work  only  until  repairs  are  made. 
In  a  wet  shaft,  on  the  other  hand  (particularly  where  the 
ordinary  types  of  sinking  pumps  are  used),  an  hour's  lack 
of  steam  may  submerge  the  pumps  and  allow  the  shaft  to 
fill.  It  will  then  be  necessary  to  get  new  pumps  and 
piping  and  to  fight  the  water  down  again  from  the  top,  and 

16 


PLANT  REQUIRED  17 

weeks  or  months  may  elapse  before  sinking  can  be  re- 
sumed. 

For  a  wet  shaft  or  for  any  shaft  deeper  than  250  or  300  ft., 
the  bricked-in  return-tubular  boiler  is  the  most  satisfactory 
type.  Such  a  boiler  burns  under  normal  firing  15  to  20 
per  cent,  less  coal  than  the  ordinary  portable  boiler.  The 
difference  in  the  coal  bill  for  100  boiler  horse-power,  with 
coal  at  $4.50  a  ton,  will,  therefore,  in  three  months,  amount 
to  $300,  which  is  about  the  cost  of  bricking  in  a  100  horse- 
power return-tubular  boiler.  The  latter  also  costs  less  for 
repairs  and  is  generally  less  trouble  than  the  portable  boiler. 

For  a  short  job  the  oil  well,  or  locomotive  type,  boiler 
is  the  best.  The  size  should  be  not  smaller  than  40  horse 
power;  60  horse-power  is  better,  as  in  the  small  sizes  the 
crown  sheet  has  such  a  shallow  covering  of  water  that  it 
is  easily  burned.  The  dome  should  be  placed  on  the  barrel 
of  the  boiler;  if  over  the  crown  sheet,  the  long  stay  bolts 
connecting  the  crown  sheet  and  the  top  of  the  dome  are  likely 
to  give  trouble.  By  utilizing  the  exhaust  from  a  compressor 
or  hoist  engine,  it  is  possible  to  force  the  locomotive  boiler 
to  make  steam  greatly  in  excess  of  its  rated  capacity,  and 
this  fact  gives  it  a  great  advantage  over  other  types. 

At  coal  shafts  the  boilers  should  be  set  far  enough  away 
to  make  it  impossible  for  a  sudden  flow  of  gas  from  the  shaft 
to  become  ignited.  They  should  always  be  placed  so  as  to 
minimize  the  cost  of  handling  coal  and  ashes.  The  ground 
at  one  end  of  the  line  of  boilers  should  be  clear  of  buildings 
or  machinery,  to  allow  of  additional  units  being  placed  as 
required. 

The  piping  should  also  be  arranged  to  permit  expansion, 
not  only  of  the  plant  as  a  whole,  but  also  the  temperature 
expansion  of  the  pipe  itself.  At  the  open  end  of  the  line  of 
boilers  the  header  should  terminate  in  a  valve,  so  that  the 
additional  boilers  can  be  coupled  on  without  shutting  down 
the  plant;  if  so  many  boilers  have  to  be  added  that  a  second 
header  is  necessary,  it  should  be  connected  with  the  first  at 
both  ends,  forming  a  steam  loop.  Stiff  connections  between 


18  PRACTICAL  SHAFT   SINKING 

the  boilers  and  the  header  are  objectionable  and  are  likely 
to  cause  leaky  joints. 

A  constant  supply  of  feedwater  must  be  assured.  Dupli- 
cate feed-pumps  or  injectors,  or  a  combination  of  the  two, 
should  be  provided,  and  the  pumps  supplying  water  from  a 
stream  to  the  supply  tank  should  also  be  in  duplicate. 

A  good  feedwater  heater  will  cut  the  coal  bill  surprisingly; 
to  be  accurate,  1  per  cent,  for  every  10  degrees  the  feedwater 
is  raised.  Assuming  the  feedwater  at  50°  F.,  originally,  an 
open  heater  with  plenty  of  exhaust  steam  will  raise  its 
temperature  to  210°  and  reduce  the  fuel  consumption  16 
per  cent.  With  coal  at  $4.50  per  ton,  a  heater  will  pay  for 


FIG.  1.  —  Ingersoll-Rand  Two-stage  Straight  line  Air  Compressor 

itself  in  two  months.  An  open  heater  as  shown  in  Fig.  1  is 
more  efficient  than  a  closed  heater  and  maintains  its  effi- 
ciency; it  has  no  tubes  to  leak  and  become  covered  with 
scale;  it  saves  the  pure  water  formed  by  the  condensed 
exhaust  steam,  and  it  is  adapted  to  various  systems  of  water 
purification.  It  must,  however,  be  used  in  connection  with 
a  good  separator  for  removing  the  oil  from  the  exhaust. 
A  satisfactory  feedwater  system  for  a  plant  containing 
several  boilers  may  be  arranged  as  follows:  Feed  all  boilers 
from  a  common  header.  Provide  regulating  valves,  in 
addition  to  regular  check-  and  boiler-stop  valves  in  con- 
nections between  header  and  boilers.  Supply  water  to 
header  with  pump  large  enough  to  feed  all  boilers  with 
piston  speed  of  50  ft.  per  minute.  Use  hard  rubber,  or 


PLANT  REQUIRED 


19 


metal,  pump  valves.  Use  an  open  heater,  placing  it  with 
base  6  ft.  above  the  pump.  As  a  reserve  provide  enough 
injectors  to  feed  the  boilers  when  the  pump  is  shut  down, 
connecting  them  into  the  feed-header.  Provide  valves 
between  each  injector  and  header,  and  between  pump  and 


FIG.  2.  —  Cochrane  Feedwater  Heater 

header.  Take  steam  connections  for  injectors  and  pump 
from  mam  steam  line.  Where  freezing  weather  is  possible, 
bun'  all  outside  water  lines. 

An  ample  power  supply  for  a  single  dry  shaft  is  100 
boiler  horse-power.  For  a  wet  shaft  the  power  required 
depends  on  the  quantity  of  water  to  be  pumped.  Three 


20  PRACTICAL  SHAFT   SINKING 

thousand  boiler  horse-power  has  been  used  for  three  very  wet 
shafts,  only  two  of  them  being  worked  at  simultaneously. 

Hoisting  Apparatus.  —  For  sinking  a  shaft  through  the 
surface  soil,  a  small  stiff-leg  derrick  is  usually  erected. 
This  makes  excavation  and  timbering  cheaper  than  if  done 
by  hand,  and  it  does  not  interfere  with  placing  the  surface 
concrete  or  add  weight  to  the  ground  around  the  shaft. 
A  derrick  with  a  40-ft.  boom  and  a  30-ft.  mast,  built  of 
12  X  12  in.  timber,  is  large  enough  for  sinking.  It  can  be 
readily  swung  by  two  men  at  the  end  of  a  10-ft.  lever  bolted 
to  the  mast.  If  any  considerable  depth  is  to  be  sunk,  this 
lever  should  be  secured  by  some  kind  of  latch  to  prevent  the 
derrick  swinging  while  the  bucket  is  in  the  shaft. 

A  double-drum  friction  engine  is  best  for  a  derrick  as  it 
enables  the  engineer  to  raise  and  lower  the  boom,  and  also, 
with  the  help  of  a  winchman,to  swing  the  derrick;  7  X  10  in. 
and  8i  X  10  in.  are  convenient  engine  sizes.  A  single-drum 
sinking  engine  may  be  used  to  advantage  on  a  derrick  with  a 
fixed  boom  swung  by  hand.  The  fall  line  sheaves  should  be 
larger  than  are  ordinarily  used  for  a  light  derrick;  never  less 
than  18  in.  outside  diameter,  preferably  24  in.  A  f-in.  rope 
will  work  over  a  24-in.  sheave  without  undue  wear. 

Although  a  small  shaft  may  be  readily  sunk  with  a  der- 
rick for  200  ft.,  it  is  better  to  put  up  a  head-frame  when  the 
surface  timbering  or  masonry  is  completed.  Sinking  head- 
frames  are  often  built  unnecessarily  large  and  heavy.  A 
head-frame  40  ft.  high  and  8  X  12  ft.  in  plan  is  large 
enough  for  a  sinking  shaft.  It  may  be  built  of  timber,  with 
8  X  8  in.  posts,  3  X  8  in.  diagonal  braces,  8  X  10  in.  caps, 
and  10  X  12  in.  sheave  timbers,  as  shown  in  Fig.  3,  or,  if 
it  is  to  be  frequently  moved,  of  steel.  If  built  of  steel, 
6  X  6  X  s-in.  angles  will  form  posts  strong  enough  to 
handle  with  safety  a  5-ton  pump.  The  sheave,  as  a  rule, 
should  not  be  smaller  in  diameter  than  the  engine  drum, 
but  a  48-in.  wheel  will  give  good  service  with  1-in.  rope. 

Two  methods  are  used  for  the  disposal  of  the  spoil 
hoisted  by  the  head-frame.  In  the  first  a  broad-gage 


PLANT   REQUIRED 


21 


track  is  extended  under  the  frame  so  that  the  rope  passes 
between  the  rails;  when  a  full  bucket  has  been  hoisted  above 
the  track,  a  truck  carrying  an  empty  bucket  is  pushed  under 
it,  the  full  bucket  is  set  on  the  truck,  and  the  empty  one 
lifted.  The  truck  is  then  pushed  away  and  the  bucket 
dumped  by  a  gallows  frame  or  other  device.  Three  buckets 
are  needed  for  this  method  so  that  one  may  be  always  in 
the  bottom. 


Back 
(Open  for ••Sftoft  T/mt>eg 


40' Sinking  Heex/fromff. 

FIG.  3.  —  Sections  of  Sinking  Head-frame 

In  the  second  and  better  method,  tipping  buckets  must 
be  used.  At  one  side  of  the  head-frame  a  chute  is  built, 
high  and  long  enough  to  discharge  spoil  into  a  dump  car, 
its  upper  end  just  clearing  the  bucket  hanging  free  on  the 
rope,  Fig.  3.  On  the  cap  above  the  chute  a  "bull  chain" 
is  hung.  The  "head-man"  stands  on  a  platform  level  with 
the  top  of  the  chute  and,  when  the  bucket  is  hoisted  within 
reach,  hooks  the  bull  chain  into  the  bail.  The  bucket  is 
lowered  slightly,  swings  out  over  the  chute,  and  is  dumped. 
The  complete  operation  may  be  performed  in  30  seconds. 


22  PRACTICAL  SHAFT  SINKING 

This  method  requires  only  two  buckets.  Three-quarter  inch 
common  chain  will  serve  for  the  bull  chain.  Its  hook 
should  be  provided  with  a  handle.  At  a  deep  shaft,  the 
bottom  of  the  chute  should  be  covered  with  pieces  of  old 
rail  laid  lengthwise,  and  in  rock  that  breaks  into  large 
lumps  a  gate  is  advisable  to  protect  the  dump  car.  A 
small  house  is  usually  built  for  the  head-man  at  the  plat- 
form level. 

To  facilitate  hooking  the  shaft  rope  to  the  bucket,  3  or 
4  ft.  of  chain  is  inserted  between  the  rope  and  the  hook. 
The  chain  should  be  welded  into  a  closed  socket  babbitted 
to  the  rope,  and  its  links  should  be  6  in.  long  so  as  to  afford 
a  good  hand  grip.  Safety  catches  are  provided'  on  the 
hook. 

Shaft  buckets  are  circular  in  plan  and  contain  from  J  to 
1  cubic  yard,  depending  on  the  depth  of  the  shaft  and  the 
size  of  the  engine.  A  flared  bucket,  Fig.  6,  discharges  rock 
more  freely  than  a  cylindrical  one;  a  convenient  size  is  3  ft. 
6  in.  top  diameter  and  2  ft.  10  in.  bottom  by  2  ft.  9  in.  high. 
The  bail  is  secured  to  the  bucket  trunnions  by  straps  and 
bolts,  so  that  it  may  be  easily  removed  for  repairs.  A 
wooden  inner  bottom  is  sometimes  used  to  cushion  the  blows 
from  pieces  of  rock.  Every  bucket  should  have  two  latches, 
and  two  lugs  to  prevent  its  dumping  in  the  wrong  direction. 

The  dump  car  that  will  not  be  knocked  to  pieces  by  large 
rocks  falling  from  the  chute  must  be  very  strongly  built. 
Its  other  qualifications  depend  on  the  nature  of  the  dump. 
Ordinarily,  where  the  dump  is  close  to  the  shaft,  and  the  car 
is  pushed  by  hand,  a  36-in.  gage,  all-around  dump  car, 
with  wheels  loose  on  the  axle,  is  best.  The  simpler  the 
construction  the  better. 

A  cheap  and  easily  erected  head-frame,  for  use  when  the 
regular  plant  is  not  available,  consists  of  a  tripod  made  of 
three  poles  bolted  together  at  the  top.  This  is  set  up  over 
the  excavation,  a  snatch  block  is  attached  at  the  top  and 
another  at  the  foot  of  one  of  the  legs,  and  small  buckets  are 
hoisted  by  a  team  of  mules,  pulled  to  one  side  and  dumped 


PLANT  REQUIRED  23 

on  a  platform  by  hand.  The  buckets  are  made  out  of  a  half 
oil  barrel,  fitted  with  extra  hoops  and  a  bail. 

For  depths  of  less  than  500  ft.,  an  81  X  10  in.  double- 
cylinder  end-friction  hoisting  engine  with  a  41 -in.  drum 
will  do  satisfactory  work.  The  friction  and  brake  levers 
are  most  convenient  if  set  in  a  stand  back  of  the  drum,  as 
is  customary  with  larger  engines.  Both  should  have  latches; 
on  the  brake  lever  a  latch  is  imperative.  An  engine  of  this 
size  will  hoist  a  loaded  bucket  weighing  2500  Ibs.  350  ft.  in 
a  minute;  a  round  trip  from  this  depth,  including  dumping 
the  bucket,  can  be  made  in  two  and  a  half  minutes. 

At  depths  greater  than  500  ft.,  the  weight  of  the  rope 
and  the  long  hoist  make  a  larger  engine  necessary.  Rever- 
sible link-motion  geared  engines  similar  to  that  shown  in 
Fig.  7  are  generally  used,  the  sizes  varying  from  10  X  12  in. 
to  14  X  20  in.  First-motion  engines  are  sometimes  used 
for  great  depths.  A  10  X  12  in.  geared  engine  running  at 
a  speed  of  400  ft.  per  minute  has  a  hoisting  capacity  of 
4500  Ibs.,  and  will  handle  muck  from  a  depth  of  800  ft. 

The  drum  should  be  grooved  so  that  the  rope  will  wind 
regularly  and  not  cut  itself.  The  size  of  rope  used  for 
sinking  runs  from  f  in.  up,  but  sizes  greater  than  1  in.  are 
unnecessary.  A  1-in.  crucible  cast-steel  rope  has  a  breaking 
strength  of  34  tons;  it  weighs  1.58  Ibs.  per  foot,  and  there- 
fore, when  hoisting  a  3000-lb.  bucket,  has  a  factor  of  safety 
of  11  at  a  depth  of  2000  ft.  This  factor  is  ample,  and  there 
is  no  use  in  consuming  power  in  hoisting  additional  weight. 

Many  lives  depend  on  the  brake  of  a  sinking  engine, 
and  it  should,  therefore,  be  made  large  and  strong  beyond 
possibility  of  fracture.  In  the  case  of  a  band  brake,  the 
diameter  of  the  part  of  the  drum  gripped  by  the  band 
should  be  as  great  or  greater  than  that  of  the  drum  itself, 
and  the  lever  should  tighten  the  band  in  the  direction  of 
the  pull  of  the  rope,  the  other  end  of  the  band  being  rigidly 
attached  to  the  frame  of  the  engine.  A  good  brake,  capable 
of  stopping  the  drum  every  time  within  an  inch  of  the  mark, 
is  not  only  a  safeguard,  but  a  great  assistance  to  sinking, 


24  PRACTICAL  SHAFT  SINKING 

especially  in  setting  up  the  bar  and  machine  or  in  handling 
timber. 

Double-drum  engines,  necessarily  friction  operated,  are 
built  for  sinking  purposes,  one  drum  being  used  for  the 
bucket,  the  other  for  handling  pumps,  piping,  etc.  The 
second  drum  introduces  another  set  of  gears,  causing  addi- 
tional friction  and  wear,  even  when  running  idle,  and  costs 
as  much  as  a  small  independent  engine,  which  is  in  every 
way  preferable.  A  compound-geared,  reversible  link-motion 
engine,  of  the  type  used  for  swinging  derricks,  makes  a  good 
engine  for  handling  pumps.  The  7  X  8  in.  size  will  hoist 
7000  Ibs.  on  a  single  rope.  The  engine  can  be  started  and 
stopped  just  where  desired,  and  there  is  no  danger  of  a  heavy 
load  getting  out  of  control. 

Electric  hoists  are  now  built  by  several  firms  in  sizes 
equivalent  to  their  standard  steam  engines,  and  operate 
satisfactorily  with  various  types  of  motors.  Gas  engines 
have  found  a  very  limited  application  to  hoisting,  but  small 
gasoline  hoists  can  be  bought. 

Signals  are  given  the  engineer  by  a  "bell"  in  the  engine 
house.  It  consists  of  a  small  whistle  or  a  hammer  striking 
a  triangle,  and  is  operated  by  a  wire  leading  to  a  bell-crank 
on  top  of  the  shaft,  thence  to  the  bottom.  A  coil  of  wire  is 
usually  clamped  to  the  horizontal  arm  of  the  bell-crank 
and  paid  out  as  the  shaft  deepens.  The  weight  of  the  wire 
in  the  shaft  is  counterbalanced  by  weights  hung  on  a  third 
arm  of  the  bell-crank  or  otherwise  arranged.  No.  6  gal- 
vanized-iron  wire  is  good  for  a  500-ft.  shaft;  for  greater 
depth  1-in.  strand  is  better.  The  bell-crank  may  be  con- 
veniently placed  in  the  head-man's  shanty. 

Regular  stopping  places  for  the  bucket,  such  as  the 
" steady,"  are  marked  by  the  engineer  by  tying  cotton  cord 
around  the  rope.  It  is  to  enable  him  to  see  these  marks 
more  readily  that  the  lever  stand  should  be  placed  behind 
the  engine. 

After  a  shaft  has  reached  a  depth  of  about  200  ft.,  it 
becomes  necessary  to  steady  the  bucket  very  carefully 


PLANT  REQUIRED  25 

before  hoisting,  to  prevent  its  striking  the  timber.  With  a 
common  rope  the  bucket  also  rotates  rapidly  on  a  long  hoist. 
To  avoid  these  effects,  guides  and  a  "billy,"  or  " dummy," 
Fig  4,  are  installed.  The  billy  is  a  light  frame  of  wood  or 
iron  composed  of  two  upright  parts  engaging  the  guides,  and 
a  cross-bar,  through  the  middle  of  which  the  rope  passes. 
It  is  carried  by  a  buffer,  clamped  to  the  rope  4  or  5  ft.  above 
the  chain  socket.  The  guides  usually  are  terminated  at 
the  bottom  of  the  last  placed  section  of  permanent  lining, 
and  buffer  blocks  stop  the  billy  at  this  point,  the  rope 
running  through  the  hole  in  the  cross-piece  as  the  bucket 
descends  into  the  bottom.  If  the  billy  is  made  of  wood, 
this  hole  should  be  lined 
with  iron  to  prevent  cut-  ' 

ting.  Old  rubber  pump 
valves  make  good  buffers  on 
the  rope  and  on  the  stop- 
blocks. 

Both  wooden  and  wire- 
rope  guides  are  used  for  the 
billy,  but  even  where  the  FIG.  4. —  "Billy" 

permanent    guide    timbers 

are  available,  rope  is  to  be  preferred.  It  can  not  only  be 
placed  more  quickly  and  cheaply  than  timber,  but  it  is 
safer.  With  timber  there  is  a  likelihood  of  the  billy  stick- 
ing, and  then  jarring  loose  and  falling  on  the  bucket;  with 
wire  rope  this  danger  is  avoided.  At  a  shaft  in  Western 
Pennsylvania  several  years  ago,  the  billy,  after  sticking 
on  some  ice  on  the  wooden  guides,  fell  and  killed  four  men. 

The  use  of  a  billy  prevents  any  rotation  of  the  bucket 
above  the  bottom  of  the  guides,  but  below  them  the  rotation 
seems  intensified.  To  obviate  this  difficulty,  non-rotating 
ropes  have  been  devised.  One  form  consists  of  a  core  and 
two  layers  of  7-wire  strands  wound  right-handed  and  left- 
handed,  respectively.  The  wires  in  the  strands  may  be 
wound  either  common  or  lang-lay.  These  ropes  fulfil  their 
purpose  (in  fact  are  specified  in  some  recent  contracts),  but 


26  PRACTICAL  SHAFT  SINKING 

do  not  wear  particularly  well.  It  is  impossible  to  use  an 
ordinary  lang-lay  rope  for  sinking  as  it  will  entirely  untwist. 

Secondary-power  Producers.  —  The  most  important  of 
the  secondary-power  producers  around  the  sinking  plant 
is  the  air  compressor.  As  yet,  electricity  has  been  unable 
to  compete  with  steam  or  compressed  air  as  a  motive  power 
for  rock  drills  or  sinking  pumps;  for  underground  work  air 
has  incidental  advantages  over  electricity  in  that  it  assists 
ventilation  and  cannot  ignite  explosive  gases. 

The  simple  straight-line  air  compressor  is  the  favorite 
for  sinking.  It  is  made  by  a  number  of  firms;  the  Ingersoll- 
Rand  Co.'s  sizes  range  from  10-in.  steam  X  10i-in.  air 
X  12-in.  stroke  to  24-in.  steam  X  24|-in.  air  X  30-in. 
stroke,  with  capacities  of  177  and  1223  cu.  ft.  of  free  air  per 
minute,  respectively.  It  is  more  efficient  mechanically 
than  most  small  engines,  and  is  wonderfully  dependable 
with  reasonable  care.  The  16  X  161  X  18-in.  size  is  a 
convenient  one  for  a  pair  of  shafts;  it  has  a  capacity  of  500 
cu.  ft.  and  will  readily  operate  four  drills  and  a  small  pump. 

A  straight-line  compressor  with  a  two-stage  air  end, 
Fig.  1,  is  made,  which,  according  to  the  statements  of  the 
manufacturers,  ought  to  be  a  good  investment.  With  a 
steam  consumption  of  45  Ibs.  per  indicated  horse-power  at 
half  cut-off,  the  simple  compressor  has,  for  each  indicated 
horse-power,  a  capacity  of  5  cu.  ft.  of  free  air  per  minute 
compressed  to  100  Ibs.  With  the  same  steam  consumption 
the  two-stage  compressor  will  deliver  15  per  cent,  more  air. 
For  500  cu.  ft.  free  air  per  minute,  the  saving  of  the  two- 
stage  over  the  simple  type  will  therefore  amount  to  15  per 
cent.  X  i  X  500  cu.  ft.  X  45  =  675  Ibs.  steam  or  150  Ibs. 
coal  per  hour.  With  the  compressor  operating  to  capacity 
twenty  hours  per  day,  six  days  a  week,  the  saving  in  three 
months,  with  coal  at  $4.50  per  ton,  would  thus  be  $525. 
This  is  somewhat  more  than  the  difference  in  cost  of 
the  two  machines. 

On  tunnels  and  similar  work,  where  a  number  of  shafts 
and  headings  are  to  be  driven  along  a  line,  it  is  economical 


PLANT   REQUIRED  27 

to  put  up  a  central  power  plant  at  a  point  where  coal  can 
be  most  conveniently  delivered  and  to  pipe  the  air  to  the 
several  openings.  For  installations  of  this  kind,  the  cross- 
compound  condensing  steam,  two-stage  air  compressor  is 
best.  A  good-sized  machine  of  this  type,  fitted  with  Corliss 
valves  and  a  well-designed  inter-cooler,  has  a  steam  con- 
sumption of  16  to  18  Ibs.  per  indicated  horse-power  hour,  and 
will  compress  5.8  cu.  ft.  of  air  per  minute  per  indicated 
horse-power.  All  the  figures  given  for  air-compressor  power 
apply  to  100  Ibs.  receiver  pressure,  the  machines  operating 
at  sea  level. 

When  the  air  is  piped  to  a  considerable  distance,  an 
after-cooler  at  the  compressor  will  condense  a  large  propor- 
tion of  the  water  vapor  carried,  and  thus  prevent  the  for- 
mation of  ice  in  the  pipes  and  valve  chests.  Freezing  is 
also  prevented  by  the  use  of  a  reheater  in  the  pipe  where 
the  air  is  to  be  used,  which  also  increases  the  power  obtain- 
able from  cold  compressed  air  at  a  small  expenditure  of 
fuel.  A  good  reheater  will  receive  air  at  60°  and  deliver  it 
at  240°,  thus  raising  the  volume  and  the  available  power  25 
per  cent,  with  an  insignificant  coal  consumption,  if  the  air 
can  be  used  before  it  cools. 

Electric  light  is  now  essential  for  effective  night  work 
anywhere,  and  is  particularly  useful  at  a  sinking  shaft  where  . 
the  outside  work  must  be  carried  on  under  all  conditions  of 
wind  and  weather,  and  the  inside  work  sometimes  under 
conditions  (such  as  great  quantities  of  falling  water  and  ex- 
plosive gases)  that  make  the  maintenance  of  an  open  flame 
impossible.  Very  little  power  is  needed,  as  two  arc  lights 
and  30  incandescents  will  give  abundant  illumination  in 
and  around  any  single  shaft.  A  5-kilowatt  generator  is 
thus  large  enough  to  light  a  pair  of  shafts,  but  as  it  may  be 
desirable  to  supply  light  to  other  work  near  the. shafts,  or 
to  run  one  or  two  small  motors,  it  is  better  to  double  this 
size.  Small  direct-connected  units  are  made  that  are 
compact  and  easily  handled  but  are  expensive;  and  the  care 
that  machinery  gets  around  construction  work  does  not 


28 


PRACTICAL  SHAFT  SINKING 


warrant  the  use  of  a  small  and  delicate  high-speed  engine. 
A  cheap  and  satisfactory  light  plant  is  formed  by  an  111 
kilowatt  generator,  belt-connected  to  an  8  X  12  in.,  hori- 
zontal, medium-speed,  automatic  engine.  The  vol 


FIG.  5.  —  Sinking  Head-frame  showing  Dumping  Arrangement 

should  not  be  higher  than  220;  even  110  will  give  quite  a 
severe  shock  to  a  man  who  is  soaking  wet. 

The  outside  wiring  calls  for  no  especial  comment.  In 
the  shaft,  however,  very  thorough  insulation  is  required 
on  account  of  the  constant  fall  of  water.  In  sinking,  the 
bottom  lights  must  be  raised  before  every  shot,  and  it  is 


PLANT  REQUIRED  29 

most  convenient  to  suspend  them  by  their  own  wire  from  a 
reel.  The  reel  may  be  kept  on  top  of  the  shaft  for  400  or 
500  ft.  of  sinking,  and  then  moved  down  to  reduce  the  weight 
of  hanging  wire.  A  suitable  reel  may  be  cheaply  built  of 
wood  by  a  carpenter.  The  two  wires  should  wind  on  sepa- 
rate drums  on  the  same  shaft,  so  that  they  will  hang  entirely 
clear  of  each  other.  If  a  wooden  shaft  is  used,  the  journals 
may  be  covered  with  copper  strips,  and  made  to  serve  as 
collecting  rings. 

Six  16  candle-power  bulbs  arranged  as  a  cluster  will  light 
the  shaft  bottom.  They  should  be  set  in  waterproof  sockets 
and  protected  against  breakage  by  wire  screens.  An 
inverted  dishpan  hung  above  the  cluster,  with  the  wires 
passing  through  a  hole  in  the  middle,  will  shed  falling  water, 
and  also  act  as  a  reflector.  The  wires  must  be  heavily 
insulated  where  they  pass  through  the  pan. 

Of  the  auxiliary  mechanisms  of  a  sinking  plant,  machine 
tools  and  small  fans  or  blowers  may  be  advantageously 
motor  driven.  If  a  large  fan  is  necessary,  it  is  simpler  and 
safer  to  drive  it  direct  by  a  separate  engine.  A  very  useful 
machine,  that  may  be  either  engine  or  motor  driven,  is  a 
swinging  cut-off  saw  for  cutting  lagging  to  length.  Such  a 
machine  will  pay  for  itself  on  a  single  deep  shaft. 

Buildings.  —  After  the  machinery  has  been  selected  and 
set  up,  it  is  necessary  to  house  it  and  the  men  that  operate 
it.  The  cheap  and  obvious  building  materials  for  temporary 
work  are  1-in.  boards  and  tar  paper.  They  are  also  highly 
inflammable,  and  on  that  account  should  be  used  with  dis- 
cretion when  it  comes  to  covering  valuable  machinery.  It 
is  surprising  how  completely  the  burning  of  a  board  shanty 
will  wreck  an  engine  inside  it.  Twenty-two  gage  corru- 
gated iron  can  be  bought  and  erected  nearly  as  cheaply  as 
boards  and  paper,  and  should  at  least  be  used  for  covering 
the  compressors  and  engines.  In  cold  weather  it  is  hard  to 
heat  a  corrugated  iron  building,  hence  boards  are  preferable 
for  shifting  shanties,  etc. 

The  buildings  needed  around  a  sinking  shaft  are:  Boiler 


30 


PRACTICAL  SHAFT  SINKING 


house;  compressor  and  dynamo  house;  engine  house;  shift 
shanty;  blacksmith  and  machine-shop;  powder  house;  oil 
house;  powder  thawing  house;  office  and  tool  house. 

The  sizes  and  styles  of  these  depend  on  the  size  of  the 
job  and  the  desires  of  the  man  who  is  running  it.  He  may 
consolidate  or  omit  some  of  them.  In  general,  however, 
they  all  have  different  functions  and  should  be  separate. 

If  the  boilers  are  under  the  same  roof  as  the  machinery, 
they  should'  be  divided  from  it  by  a  tight  partition  to  keep 
cinders  and  dirt  out  of  the  bearings.  The  shift  shanty 
should  be  adjacent  to  the  shaft,  large  enough  for  all  the  men 


lays  and  Latches  on 
Both  Sides  of  Bucket 


J"Jf£  'Trunnion  ftmf. 
/feiywS  Bottom 


Detar/  o/  Connect  ion 
o/ Ba/J  to  Trunnion. 

FIG.  6.  —  Details  of  Sinking  Bucket 

on  the  shift  to  change  their  clothes  at  once,  and  should  have 
plenty  of  pegs  for  drying  clothes,  and  a  good  stove  or  radiator. 
The  powder  house  and  the  oil  house  should  be  separated 
from  each  other  and  the  former  placed  some  distance  from 
the  job.  They  should  be  only  large  enough  to  contain  the 
stocks  of  dynamite  and  oil  actually  needed,  and  should 
be  built  of  iron  to  lessen  the  risk  of  fire  and  lighting. 
Caps  and  exploders  should  never  be  stored  with  dynamite. 
The  powder  thawing  house  is  preferably  a  box  or  closet  that 
will  hold  three  or  four  boxes  of  dynamite,  and  is  heated  by 
steam  coils.  It  should  be  so  constructed  that  loose  sticks 
of  powder  cannot  come  in  contact  with  the  hot  pipes. 
Thawing  boxes  covered  with  manure  are  sometimes  used, 
but  are  not  safe,  as  manure  is  liable  to  spontaneous  com- 
bustion. 


PLANT  REQUIRED  31 

The  blacksmith  and  machine  shop  should  contain  a 
forge  fitted  with  bellows- or  hand  blower  as  well  as  a  blast 
connection  to  the  air  line,  benches  with  common  and  pipe 
vises,  a  grindstone,  a  small  drill  press,  and,  on  a  good-sized 
job,  a  pipe  cutting  and  threading  machine.  The  small  tools 
should  comprise  blacksmith  and  drill-sharpening  tools,  pipe 
dies  and  cutters,  bolt  dies  and  taps,  a  ratchet  drill,  hacksaw, 
hammers,  monkey  and  pipe  wrenches,  chain  tongs,  etc. 
A  good  assortment  of  miscellaneous  pipe  fittings  and  drill 
repairs  should  be  kept  on  hand. 


FIG.  7.  —  Double  Spur  Gear  Reversible-link  Motion  Lidgerwood  Hoist 

The  contents  of  the  tool  room  also  depend  on  the  size  of 
the  job,  but  a  good  equipment  saves  money  in  the  end. 
The  following  articles  are  either  necessary  or  very  useful: 
Good  assortment  of  round  and  flat  blacksmith  iron,  assorted 
nuts  and  washers,  packing  for  engine  and  compressor 
glands  and  pumps,  gasket,  waste,  oil  cans,  torches,  crosscut 
saw,  crowbars,  striking  hammers,  picks  and  shovels,  assorted 
nails,  manila  rope  and  blocks,  cant  hooks,  lever  jacks,  etc. 

The  general  layout  of  the  job  depends  so  largely  on  local 


32  PRACTICAL  SHAFT  SINKING 

recommendations.  The  location  of  the  boilers  has  already 
been  discussed;  if  a  railroad  siding  leads  to  the  shaft,  it  is 
well  to  place  the  line  of  boilers  parallel  to  it,  so  that  coal  can 
be  unloaded  directly  into  bunkers.  The  storage  and  sub- 
sequent handling  of  timber  must  also  be  considered.  The 
temporary  buildings  should  be  so  located  that  they  will  not 
be  put  into  a  hole  by  the  encroachment  of  the  dump;  the 
position  of  the  dump  itself  must  be  considered  with  relation 
to  the  drainage  of  the  surrounding  ground.  The  sinking 
engine,  boilers,  and  machinery  (as  is  usually  specified) 
should  not  interfere  with  the  erection  of  the  permanent 
mining  plant.  Lastly,  it  may  be  again  stated  that  too 
much  attention  cannot  be  given  to  the  piping  system  all 
over  the  job  as  regards  tightness,  drainage,  and  insulation. 

Cost.  —  The  cost  of  a  plant  for  a  single  shaft,  assuming 
a  depth  of  about  500  ft.  and  a  moderate  inflow  of  water, 
say  30  or  40  gallons  a  minute,  is  as  follows: 

Sinking  engine    $1,000 

Two  80  horse-power  boilers  and  setting 1,800 

Pipe  and  auxiliaries  500 

150  horse-power  heater 300 

14-inch  compressor 1,750 

Three  drills  and  steel 1,000 

Shaft  bar  and  clamps    100 

Derrick 400 

Head-frame 500 

Two  buckets  150 

Rope 150 

Buildings 500 

Dump  cars  and  rail 300 

Electric  plant,  10  kilowatts   750 

Two  pumps 500 

Small  tools 500 

Total $10,200 

These  figures  are  based  on  the  cost  of  new  machinery, 
and  are  large  enough  to  include  the  necessary  accessories. 
The  cost  of  erecting  and  dismantling  such  a  plant  will  be 
from  $1000  to  $2000,  depending  on  location,  labor  condi- 
tions, etc. 


CHAPTER   III 

SINKING  THROUGH  SURFACE  —  SOFT  GROUND  —  WOODEN 
SHEETING  —  STEEL  SHEETING  —  CAISSONS  OF  STEEL, 
WOOD,  OR  CONCRETE. 

SINKING  THROUGH  SURFACE 

IN  most  localities  a  certain  amount  of  soil  or  soft  ground 
overlies  the  ledge  rock.  Its  depth  varies  from  nothing  to 
hundreds  or  even  a  thousand  feet,  and  its  nature  is  as  varied 
as  that  of  the  rocks  which  it  covers.  The  shaft  sinker  is 
interested  chiefly  in  its  consistency,  which  determines 
whether  the  penetration  of  the  surface  will  be  the  easiest  or 
the  most  arduous  and  expensive  part  of  his  job. 

There  is  no  hard  and  fast  line  of  demarcation  between 
firm  ground  and  running  ground;  every  degree  of  hardness 
or  softness  can  be  found  from  boulder  clay  to  river  silt, 
but  ordinarily  in  sinking,  ground  is  considered  firm  when  the 
excavation  can  be  carried  ahead  of  the  support,  and  soft 
when  the  support  must  be  driven  ahead  of  the  excavation. 
In  the  first  classification  are  included  boulder  clay,  ordinary 
dry  blue  or  yellow  clay,  cemented  or  clayey  gravel  and  most 
loam  soil;  loose  sand  and  gravel  and  silt  come  under  the 
second. 

The  amount  of  water  in  the  ground  has  a  very  great 
effect  on  its  firmness,  as  is  shown  by  the  caving  of  excava- 
tions after  a  rain  storm;  conversely,  soft  wet  ground  may  be 
made  comparatively  firm  by  removing  the  water.  The 
commonest  application  of  this  principle  is  the  use  of  com- 
pressed air  for  driving  quicksand  tunnels  without  the  use 
of  a  shield.  In  this  case  the  water  is  forced  back  away  from 
the  face  into  the  surrounding  ground,  and  timbering  opera- 


34 


PRACTICAL  SHAFT   SINKING 


tions  can  be  performed  readily,  which  would  be  utterly 
impossible  if  the  water  were  allowed  to  flow  into  the  bore 
of  the  tunnel. 

A  trench  in  quicksand  was  recently  driven  at  Gary,  Ind., 
with  very  great  success;  here  the  water  was  drained  in 
advance  of  the  excavation  through  a  number  of  small  per- 
forated pipes  driven  into  the  ground  and  connected  at  the 


FIG.  8.  —  Hanging  Timbers  in  Firm  Earth 

upper  ends  to  the  suction  side  of  a  pump.  The  writer  knows 
of  no  case  in  which  quicksand  has  been  drained  in  advance 
of  the  excavation  of  a  sinking  shaft  by  driving  small  suction 
pipes  into  it,  but  in  view  of  the  success  of  the  plan  in  the 
trench  instanced  above,  he  sees  no  reason  why  it  could  not 
be  worked  out  for  a  shaft. 

Compressed  air  is  extensively  •  used  for  sinking  bridge 
piers  and  other  caissons  through  soft  ground;  the  applica- 
tion of  this  process  to  shafts  will  be  considered  later. 


SINKING  THROUGH   SURFACE  35 

Concrete  masonry  is  now  almost  entirely  used  for  per- 
manently supporting  the  sides  of  shafts  in  soft  ground. 

The  methods  in  use  for  temporary  support  while  sinking 
are: 

Timbering;  this  heading  includes  the  driving  of  wooden 
or  steel  sheet  piling,  as  well  as  forepoling. 

Caissons;  these  may  be  open  drums,  or  closed  drums  sunk 
under  compressed  air. 

Iron  sinking  drums  and  shoes,  forced  down  hydraulically. 

The  freezing  process. 

The  first  method  is  applicable  to  comparatively  easy 
surface  conditions;  the  others  to  more  difficult  conditions. 
The  last  two  have  been  developed  in  Europe  for  sinking 
through  great  depths  of  sand  or  mud,  and  are  not  extensively 
used  in  this  country.  The  various  methods  will  be  treated 
in  order. 

Timbering  —  While  sinking  through  ordinary  surface 
ground  the  sides  of  the  shaft  are  usually  supported  by 
square-framed  horizontal  sets  of  timbers  with  vertical 
lagging  behind  them.  The  distance  between  the  sets 
depends  upon  the  firmness  of  the  ground,  and  varies  from 
about  6  ft.  as  a  maximum  to  nothing  for  "skin  to  skin" 
timbers  in  soft  material.  In  square-framed  sets  the  end 
and  side  pieces  are  termed  "end  plates"  and  "wall  plates," 
respectively;  the  cross-struts,  "buntons,"  and  the  posts 
which  separate  the  sets,  "punch  blocks." 

FIRM    GROUND 

The  cheapest  kind  of  sinking  is  afforded  by  earth  that 
does  not  require  blasting,  yet  is  stiff  enough  to  stand 'ver- 
tically for  4  or  5  ft.  without  support.  In  such  material  the 
usual  procedure  is  to  commence  the  excavation  just  large 
enough  to  admit  the  timber  and  lagging,  and  to  carry  the 
sides  down  vertically  without  support  as  far  as  it  is  safe  to 
do  so.  The  timbering  is  then  started  on  the  bottom  and 
brought  up  to  the  surface  of  the  ground.  Two  or  more 
heavy  bearing  timbers,  long  enough  to  extend  4  or  5  ft. 


36  PRACTICAL  SHAFT  SINKING 

beyond  the  lagging  at  each  end,  are  laid  across  the  shaft  on 
the  surface  and  their  ends  are  supported  by  blocking  them 
solidly  against  the  ground,  Fig.  8.  The  sets  of  timber  are 
then  hung  from  these  bearing  timbers  by  heavy  rods,  and 
sinking  is  resumed.  As  soon  as  4  or  5  ft.  of  ground  is 
removed,  another  set  is  placed  on  the  bottom,  hung  with 
rods  to  the  set  above,  and  the  lagging  is  worked  in  back  of 
them  in  pieces  just  long  enough  to  bear  on  both  sets.  The 
process  is  then  repeated.  Bearing  timbers  are  usually 
placed  over  the  end  plates  and  over  each  row  of  buntons, 
and  punch  blocks  are  set  at  the  corners  and  under  the  ends 
of  all  buntons.-  10  X  12  in.  timber  sets,  spaced  4  ft.  center 
to  center  and  braced  so  that  the  longest  span  will  not  exceed 
12  ft.,  will  safely  support  firm  earth  for  a  depth  of  60  or  70  ft. 
The  weight  of  the  timbers  is  partly  carried  by  the  friction 
of  the  earth  against  the  lagging,  and  the  hanging  bolts  are 
not  subjected  to  great  stress;  they  may  sometimes  be  en- 
tirely omitted.  A  ledge  of  earth  is  in  this  case  left  under 
the  bottom  set,  and  the  middle  of  the  shaft  excavated; 
inclined  posts  are  then  wedged  between  the  shaft  bottom 
and  the  timbers  and  the  ledge  removed.  It  is  generally 
safer  to  use  1|  or  IJ-in.  hanging  bolts,  however.  Exca- 
vations of  this  character  can  be  done  for  a  total  labor  cost, 
including  the  placing  of  timber,  of  $1.50  to  $2  per  cubic  yard. 
As  the  softness  of  the  ground  increases,  the  distance  between 
sets -is  decreased.  Sometimes  the  lagging  is  omitted  and 
the  lower  set  worked  in  immediately  under  the  one  above. 
The  consideration  of  this  plan  properly  belongs  under  soft- 
ground  work. 

SOFT   GROUND 

Wooden  Sheeting.  —  When  ground  is  so  soft  that  it  will 
not  stand  vertically  at  all,  it  becomes  necessary  to  support 
it  in  advance  of  the  excavation.  The  commonest  method 
of  doing  this  in  any  kind  of  pit  is  to  enclose  the  area  to  be 
dug  out  with  a  coffer  of  sheet  piling,  driven  by  hand  or 
power,  Fig.  9,  and  to  brace  the  inside  of  the  coffer  as  the 


SINKING   THROUGH   SURFACE 


material  is  removed.  In  starting  a  shaft,  two  sets  of  timber, 
one  5  ft.  or  so  above  the  other,  are  set  up  as  a  guide  frame, 
and  the  sheeting  driven  around  them.  The  top  soil  is 
usually  firm  enough  to  enable  these  sets  to  be  placed  below 
the  surface,  but  this  is  not,  of  course,  essential.  If  the  sets 
are  placed  above  the  surface,  outside  waling  pieces  are 
bolted  through  the  sheeting  at  the  top  set  in  order  to  hold 
the  top  of  the  sheeting  in  line. 


•So// 


FIG.  9.  —  Successive  Courses  of  Sheeting 

In  dry  sand  or  other  loose  ground  that  does  not  contain 
much  water,  the  sheeting  is  driven  as  the  excavation  pro- 
gresses, and  the  points  of  the  piles  are  kept  only  slightly 
below  the  bottom.  Two-inch  planks  in  12  to  16  ft.  lengths 
are  commonly  used  in  this  case  and  are  driven  by  hand  with 
heavy  wooden  mauls.  The  heads  should  have  beveled 
edges  to  prevent  splitting,  Fig.  10,  and  for  hard  driving,  or 
with  soft  wood,  a  plate-iron  cap  may  be  used  to  advantage. 
By  thus  protecting  their  heads,  the  planks  can  be  driven  to 
their  full  length;  a  second  course  of  sheeting  is  then  driven 
inside  the  timbering  of  the  first  course,  and  so  on  until  the 
required  depth  is  reached.  The  economical  limit  for  this 


38 


PRACTICAL   SHAFT  SINKING 


method,  however,  is  about  50  ft.,  as  it  necessitates  starting 
the  shaft  much  larger  than  the  minimum  required  size; 
some  additional  allowance  must  be  made  on  every  course 
for  possible  distortion  and  for  inward  bending  of  the  sheeting 
at  the  points.  Let  us  assume  50  ft.  of  surface,  10  X  10  in. 
timber,  and  2-in.  lagging,  and  a  shaft  12  X  24  ft.  with  a 
4-ft.  concrete  curb  wall.  Four  courses  of  sheeting  will  be 
required,  the  last  20  ft.  4  in.  X  32  ft.  4  in.  outside,  its  wall 
plates  to  be  buried  in  the  concrete.  Allowing  6  in.  all 
around  each  time  for  distortion  or  squeezing,  the  third  set 
will  be  23  ft.  4  in.  X  35  ft.  4  in.,  the  second  26  ft.  4  in.  X 


FIG.  10 


FIG.  11 


38  ft.  4  in.,  and  the  first  29  ft.  4  in.  X  41  ft.  4  in.  The  total 
excavation  will  thus  be  40  per  cent. .in  excess  of  that  theoreti- 
cally required  for  the  curb  wall.  Any  additional  depth 
necessitating  another  course  of  sheeting  will  increase  the 
percentage  of  useless  excavation,  and  will  require  a  larger 
quantity  of  heavy  timber. 

When  many  light  sheet  piles  are  to  be  driven,  the  work 
can  be  done  more  cheaply  with  some  form  of  power  driver 
than  with  mauls.  A  driver  like  an  enlarged  rock  drill, 
Fig.  12,  has  been  devised  for  this  purpose,  and  a  common 
drill  fitted  with  a  hammer  instead  of  a  bit  will  drive  light 
sheeting  satisfactorily.  Either  machine  is  suspended  with 
blocks  and  falls  from  a  trolley  or  tripod  over  the  line  of 
sheeting. 


SINKING  THROUGH  SURFACE  39 

Steel  Sheeting.  —  In  quicksand  or  other  wet  running 
ground,  sheeting  must  have  joints  that  are  almost  water- 
tight, and  as  it  is  impossible  to  drive  common  plank  close 
enough  to  make  a  satisfactory  coffer  in  such  material  with- 
out caulking,  some  form  of  interlocking  piling  should  be 
used.  Formerly  tongued  and  grooved,  splined,  or  Wakefield 
piling,  Fig.  11,  were  the  only  forms  available,  but  now  they 


FIG.  12.  —  Sheet-pile  Driver 

have  been  superseded  for  difficult  work  by  the  interlocking 
steel  sheet  pile.  A  slight  obstruction  will  cause  wooden 
piling  to  separate  at  the  bottom,  whereas  it  is  almost  im- 
possible to  pull  the  steel  piles  apart.  Steel  piles,  moreover, 
can  be  more  easily  driven,  will  penetrate  most  obstructions, 
and  can  be  readily  pulled  and  redriven.  There  are  a  dozen 
types  on  the  market,  each  with  its  advocates,  but  the  sim- 
plest shapes  are  those  rolled  by  the  Carnegie  Steel  Co., 
(a)  Fig.  13,  and  by  the  Lackawanna  Steel  Co.,  (6)  Fig.  13. 


40  PRACTICAL  SHAFT  SINKING 

Both  are  strong  and  satisfactory  and,  though  not  water- 
tight when  first  driven,  will  soon  become  so  in  most  ground. 

An  additional  advantage  that  steel  piles  possess  is  that 
they  can  be  obtained  in  lengths  up  to  60  ft.  and  can  be 
completely  driven  before  excavation  is  started.  When  the 
ground  is  very  bad,  they  should  be  made  to  reach  rock  so  as 
to  prevent  material  from  flowing  under  their  points.  In 
one  case  a  hole  36  X  27  ft.  6  in.  in  plan  and  27  ft.  deep  was 
needed  for  a  furnace  pit;  the  material  was  soft  quicksand 
and  rock  lay  at  an  unknown  depth.  Steel  sheet  piling  48  ft. 
long  was  obtained  and  successfully  driven  entirely  around 
the  pit  and  followed  down  4  ft.  below  the  surface.  The  first 
20  ft.  of  sand  was  easily  removed,  but  as  the  depth  increased, 
sand  began  to  flow  in  under  the  piling  and  gradually  bent 
their  points  inward,  throwing  a  terrific  strain  on  the  lowest 
set  of  timber.  A  complete  wreck  was  finally  prevented  by 
filling  the  hole  with  sacks  of  concrete  which  sank  into  the 
sand  and  supported  the  lower  end  of  the  piling.  This  en- 
abled the  desired  depth  to  be  reached,  but  it  would  have 
been  practically  impossible  to  reach  rock  if  the  hole  had  been 
intended  for  a  shaft. 

Steel  piles  and  heavy  wooden  sheet  piles  must,  of  course, 
be  driven  by  machinery.  While  a  discussion  of  pile  driving 
would  be  out  of  place  in  an  article  on  shafts,  it  may  be  said 
that,  in  the  writer's  opinion,  a  steam  hammer  is  preferable 
to  a  drop  hammer  for  sheet  piling,  whether  used  in  regular 
or  suspended  leads.  Sometimes  a  water  jet  is  necessary; 
with  a  jet  piles  can  be  easily  sunk  by  the  weight  of  the  ham- 
mer through  sands  into  which  they  cannot  be  driven 
at  all. 

The  chief  trouble  in  driving  a  steel-pile  coffer  is  in  making 
a  good  closure.  Sometimes  the  last  pile  exactly  fills  the 
gap,  but  more  often  a  lap  joint  is  made  which  is  caulked  with 
hay  or  junk.  With  care  a  good  joint  can  be  made  in  this 
way.  Steel  piling  in  short  lengths  is  used  for  cutting  off 
thin  strata  of  quicksand  encountered  some  distance  below 
the  surface.  In  this  case  the  piles  are  locked  together  all 


SINKING   THROUGH  SURFACE  41 

the  way  around  and  each  is  driven  only  2  or  3  ft.  at  a  time 
until  all  reach  the  rock. 

Steel  piling  can  be  driven  through  logs,  strata  of  cemented 
gravel,  etc.,  but  in  ground  containing  large  hard  boulders 
some  other  method  must  be  used.  At  a  quicksand  shaft  in 
Michigan,  boulders  were  encountered,  but  steel  piles  were 
driven  until  they  had  apparently  reached  the  desired  depth. 
The  coffer,  however,  could  not  be  excavated,  as  the  sand 
in  some  way  flowed  in  as  fast  as  it  was  removed.  Com- 
pressed air  was  finally  applied,  and  when  the  points  of  the 
piles  were  reached,  they  were  found  to  be  torn  apart  by  the 
boulders.  Several  piles,  bent  through  a  full  half  circle, 
were  pointing  up  the  shaft. 


(a) 

FIG.  13.  —  Sheet  Steel  Piling 

Forepoling.  —  Forepoling  was  formerly  used  for  shafts 
in  soft  ground  of  any  nature,  and  depths  of  100  ft.  have  been 
reached  in  the  worst  kind  of  material.  Under  such  condi- 
tions forepoling  is  very  slow  and  expensive,  and  although  it 
has  been  largely,  if  not  entirely,  superseded  by  the  steel 
sheet  pile  or  caisson  methods,  some  discussion  of  it  is  of 
interest.  Forepoling  is  still  widely  used  for  soft-ground 
tunnels. 

In  starting  a  shaft  which  is  to  be  forepoled  through 
difficult  ground,  strong  trusses  are  used  for  bearing  timbers 
and  the  ring  timbers  are  suspended  from  them  by  heavy 
bolts.  The  trusses  span  the  shaft,  Fig.  14,  and  their  ends 
are  supported  by  broad  cribs  or  piers  set  well  back  from 
the  edge  of  the  shaft.  They  are  built  strong  enough  to 
carry  the  weight  of  all  the  surface  timbering  and  also  of 
the  head-frame,  if  one  is  used.  After  the  head-frame,  or 
derrick,  is  ready,  digging  is  started  and  the  sides  of  the  shaft 
are  supported  by  short  piles  or  poling  boards  driven  on  a 


42  PRACTICAL  SHAFT  SINKING 

slant  so  that  they  bear  against  .the  outer  face  of  the  bottom 
set  of  timbers  and  the  inner  face  of  the  one  above.  The 
poling  boards  are  made  twice  as  long  as  the  distance 
between  the  sets  (or  longer),  so  when  one  course  is  driven 
home,  enough  ground  can  be  dug  out  to  enable  another  set 
of  timbers  to  be  placed.  The  worse  the  ground,  the  longer 
must  the  poling  boards  be  made  to  prevent  it  flowing  under 
them.  The  most  troublesome  places  are  the  corners  where 
the  boards  are  divergent,  and  the  spaces  back  of  the  buntons 
where  a  board  is  necessarily  omitted.  These  openings  are 
closed  by  short  transverse  boards  placed  as  the  excavation 
proceeds. 

After  a  depth  of  about  40  ft.  has  been  reached,  the 
pressure  of  running  ground  becomes  so  great  that  single 
sets  of  timber,  spaced  so  the  poling  boards  can  be  driven 
between  them,  will  not  support  it.  Two  or  more  timbers 
must  then  be  placed  "skin  to  skin"  to  form  the  wall  and 
end  plates,  and  as  the  depth  increases  the  spacing  of  these 
compound  sets  must  be  reduced  until  the  poling  boards 
have  to  be  driven  nearly  horizontal,  and  therefore  fail  to 
prevent  the  ground  from  rising  in  the  shaft.  This  is  where 
troubles  with  the  forepoling  method  really  begin.  Every 
inrush  of  material  causes  a  settlement  of  the  ground  around 
the  shaft,  throws  the  shaft  itself  out  of  line,  and  puts  very 
great  stresses  on  the  timbers  and  the  hanging  bolts.  In 
some  cases  heavy  timbers  driven  with  a  ram  have  been 
used  for  poling  boards.  They  were  thus  driven  deep 
enough,  and  the  successive  courses  were  separated  suffi- 
ciently, to  permit  very  heavy  ring  timbers. 

Among  the  other  plans  that  have  been  devised  to  keep 
the  bottom  down  may  be  mentioned : 

Drainage  of  the  ground  ahead  of  the  excavation  by  means 
of  a  perforated  iron  drum,  jacked  down  and  used  as  a  sump 
for  pump  suctions.  A  short  stoppage  of  the  pumps  will 
allow  the  ground  to  become  saturated  again  and  start  a 
run,  and  besides  it  is  very  hard  to  keep  pumps  running 
steadily  with  sandy  water. 


SINKING  THROUGH  SURFACE  43 

Drainage  of  the  ground  by  means  of  a  timbered  sump 
combined  with  a  system  of  floor  boards  similar  to  the  breast 
boards  used  in  soft  ground  tunnels.  This  method  is  very 
slow  and  laborious. 

Sinking  quantities  of  hay  and  brush  into  the  ground 
around  the  shaft  by  loading  them  with  pig  iron.  This 
stiffens  the  ground  to  some  extent  and  tends  to  prevent 
runs.  It  is  a  help  with  any  style  of  timbering  in  quicksand, 
and  may  even  be  necessary  in  sinking  a  caisson. 

Caissons.  —  In  the  last  ten  years  many  American  engi- 
neers have  adopted  the  sinking  drum  or  open  caisson  for 
penetrating  soft  ground.  A  hollow  cylinder  of  masonry  is 
constructed  on  the  surface  with  its  axis  vertical  and  its 
walls  tapered  outward  at  the  bottom  to  a  cutting  edge. 
The  outer  surfaces  of  the  walls  should  be  smooth  and  verti- 
cal, and  the  cutting  edge  should  be  slightly  larger  than  the 
rest  of  the  cylinder.  After  the  masonry  has  hardened  the 
earth  is  excavated  on  the  inside,  and  the  caisson  sinks  of  its 
own  weight.  It  is  kept  plumb  by  digging  out  under  the  high 
side.  When  the  top  reaches  ground  level,  another  section 
is  added  and  this  continues  until  the  cutting  edge  reaches 
rock.  This  method  has  long  been  employed  in  Germany, 
the  caissons  being  constructed  of  brick  or  stone;  in  this 
country  timber  caissons  have  occasionally  been  used.  The 
low  tensile  strength  of  brick  or  stone  and  the  difficulty  of 
sinking  wood  in  bad  ground  made  these  materials  unsatis- 
factory. Concrete,  combining  weight  with  strength,  is 
almost  ideal,  and  as  it  is  not  only  better  than  timber  and 
brick,  but  also  cheaper,  it  has  displaced  them  both  for 
building  caissons. 

Rectangular  Caissons.  —  Caissons  are  ordinarily  circu- 
lar, but  sometimes  are  made  rectangular  of  reinforced 
concrete.  A  noteworthy  example  of  this  type  is  a  shaft 
recently  sunk  for  the  D.,  L.  &  W.  R.  R.,  on  the  flats  opposite 
Wilkesbarre,  Pa.  This  shaft,  48  ft.  10  in.  X  14  ft.  in  the 
clear,  was  sunk  through  very  wet  quicksand  to  a  depth  of 
70  ft.,  in  four  months,  including  the  time  lost  in  sealing  the 


44  PRACTICAL  SHAFT  SINKING 

caisson  to  the  rock.  It  thus  affords  a  decided  contrast  to 
the  Pettibone  shaft  near  by,  which,  sunk  by  forepoling,  took 
eight  years.  Cost  figures  are  unobtainable  but  seem  unnec- 
essary. The  walls  of  the  D.,  L.  &  W.  caisson  are  5  ft.  4  in. 
thick  at  the  bottom,  2  ft.  8  in.  at  the  top,  and  are  plumb  on 
the  outside.  Two  reenforced  cross-walls  serve  as  buntons 
and  also  support  the  side  walls. 

Several  rectangular  caissons  have  been  sunk  along  the 
Monongahela  River  in  the  flats  above  Brownsville,  Pa. 
The  ground  is  not  very  bad,  but  contains  enough  soft  clay 
and  quicksand  to  make  timbering  very  difficult.  Two  of 
these  were  coal  shafts  and  were  sunk  through  50  ft.  of  sur- 
face in  two  months  in  the  winter  of  1908-09. 

Circular  Caissons.  —  A  circular  caisson  was  sunk  in  the 
autumn  of  1908  for  Shaft  No.  2  on  the  Rondout  Siphon 
of  the  Catskill  Aqueduct.  The  surface  was  about  60  ft. 
thick,  of  which  6  ft.  was  sandy  loam.  The  balance  was  a 
wet  material  that  resembled  blue  clay  when  dried  out,  but 
which  in  the  ground  was  completely  saturated.  It  flowed 
slowly  like  cold  molasses,  and  was  very  sticky.  Overlying  the 
rock  and  entirely  surrounded  by  the  muck  were  quantities 
of  hard  boulders  of  all  sizes,  which  had  to  be  blasted  from 
under  the  cutting  edge  of  the  caisson.  The  combination 
made  as  difficult  ground  to  sink  through  as  can  well  be 
conceived. 

The  shaft  desired  was  a  three-compartment  shaft, 
10  X  22  ft.  outside  the  timbers,  with  two  hoistways.  The 
caisson  was  made  21  ft.  inside  diameter,  Fig.  15,  giving 
ample  room  for  the  hoistways  and  a  ladderway;  the  area  for 
air  is  of  course  in  excess  of  that  afforded  by  the  rectangular 
shaft.  This  caisson  was  built  and  sunk  to  rock  in  two 
months,  and  a  description  of  the  method  used  for  it  will 
give  a  good  idea  of  the  process  in  general : 

It  was  decided  that  a  caisson  with  30-in.  walls  would  be 
strong  enough  and  heavy  enough  to  sink  to  rock,  and  a  steel 
shoe  or  cutting  edge  26  ft.  in  outside  diameter  was  obtained. 
This  shoe  was  formed  of  two  \  X  20  in.  plates  riveted 


SINKING   THROUGH   SURFACE 


45 


together  at  the  bottom  and  flared  at  the  top  to  include  the 
lower  part  of  the  concrete  wall.  It  was  anchored  to  the 
concrete  by  about  eighty  f  in.  X  8  ft.  countersunk-head 
bolts. 


Short  Horijonfal  Boa/zfc 
•    fo  c/dse  corner^ 


FIG.  14.  —  Forepoling 

The  shaft  site  wras  leveled,  the  shoe  assembled  upon  short 
planks  laid  on  the  ground,  and  forms  for  the  concrete  were 
started.  The  forms  were  built  of  vertical  2-in.  lagging  in 
4-ft.  lengths,  supported  inside  and  out  by  rings  of  4  X  3  in. 
angles  tied  together  by  f -in.  rods.  Five  feet  of  1:2:5  con- 
crete was  placed  and  allowed  to  set  for  a  week  to  obviate 
the  possibility  of  settlement  cracks  above  the  cutting  edge; 
10  ft.  more  was  then  placed  and,  as  soon  as  it  had  hardened 
sufficiently  to  permit  of  the  forms  being  removed,  sinking 
was  commenced.  The  mud  was  loaded  into  shaft  buckets 


46  PRACTICAL  SHAFT   SINKING 

by  men  standing  upon  plank  rafts  and  was  hoisted  with  a 
derrick.  Sometimes  the  mud  had  to  be  bailed  with  water 
buckets,  but  usually  shovels  could  be  used. 

When  the  top  of  the  first  15  ft.  of  concrete  reached  the 
ground  level,  10  ft.  more  were  added  and  excavation  com- 
menced again.  Thus  far  the  cutting  edge  had  been  very 
little  in  advance  of  the  excavation,  but  at  this  point  the  cais- 
son suddenly  dropped  7  ft.  and  the  mud  inside  rose  12  ft. 
The  cutting  edge  was  seen  no  more  until  it  had  almost 
reached  rock.  After  the  drop  20  ft.  of  concrete  were  added 
before  excavation  was  started.  Before  this  had  been  sunk 
to  ground  level,  a  stratum  of  very  soft  mud  was  encountered 
which  ran  in  under  the  shoe  and  caused  the  surface  to  cave 
on  the  side  next  the  derrick.  The  caisson  gradually  leaned 
toward  the  caving  ground  until  it  was  nearly  2  ft.  out  of 
plumb.  Sinking  was  then  stopped,  10  ft.  of  concrete  added, 
a  trench  dug  through  the  6  ft.  of  surface  clay  on  the  high 
side  of  the  caisson,  and  the  dirt  banked  against  the  caisson 
over  the  cave-in.  When  sinking  was  started  the  caisson 
began  to  right  itself.  It  soon  stopped  moving,  how- 
ever, and  the  cutting  edge  was  found  to  be  resting  on 
large  boulders  which  had  to  be  broken  with  dynamite. 
In  the  meantime  the  mud  ran  in  almost  as  fast  as  it  could 
be  hoisted  out,  and  the  caving  continued.  When  the 
cutting  edge  was  within  about  6  ft.  of  rock,  the  caisson 
literally  stuck  in  the  mud  and  refused  to  move  even  when 
the  boulders  were  blasted  out  all  around.  A  heavy  timber 
platform  was  then  built  on  top  of  the  concrete  and  loaded 
with  200  tons  of  clay.  As  the  caisson  still  stuck,  the  sur- 
rounding mud  was  agitated  by  blowing  compressed  air  into 
it  through  U-in.  pipes  which  had  previously  been  built 
into  the  wall;  a  l|-in.  pipe  was  also  used  as  a  jet  and  worked 
down  to  its  full  length  on  the  outside.  This  was  done  over 
and  over  again  all  the  way  around.  After  a  few  hours  the 
drum  started  to  sink  and  reached  rock  without  further 
trouble. 

The  trouble  in  this  case  was  due  to  the  great  stickiness 


SINKING   THROUGH  SURFACE 


47 


of  the  mud.  An  additional  thickness  of  6  in.  in  the  walls 
would  have  probably  caused  the  caisson  to  sink  without 
delay.  As  it  was,  only  forty-eight  days  elapsed  from  the 
erection  of  the  shoe  to  the  commencement  of  rock  excava- 
tion, an  average  progress  of  1.2  ft.  per  day. 


'  i 

,  p 
1 

.!': 

i  • 

'  j: 

T 
j 
i 

i 

--/"Vert  icaf 
Reinforcement 

-80-jxd'o" 
Anchor  Pods. 

.      (2-s"x£0~  P/ate 
'&><*<  F///er 

' 

1 

Tly 

4 

/  1 

/-'-p 

E 

1 

FIG.  1.5.  —  Circular  Concrete  Caisson 


Fig.  16  shows  the  shoe  and  the  lower  part  of  the  form, 
Fig.  17  shows  the  derrick,  mixer,  and  general  layout,  and 
Fig.  18  shows  the  dump  composed  of  mud  spread  out  over 
an  acre  or  more  of  ground. 

All  caissons  should  have  some  vertical  reenforcement,  so 
that  if  the  upper  part  sticks  the  lower  part  cannot  drop 
away  from  it.  The  absence  of  this  has  caused  several 
wrecks. 


48 


PRACTICAL  SHAFT  SINKING 


Steel  shoes  are  only  necessary  in  ground  containing 
boulders.  A  concrete  cutting  edge  properly  reenforced 
is  strong  enough  to  penetrate  sand  or  clay  with  safety. 

A  number  of  points  must  be  considered  in  deciding 
whether  to  use  a  rectangular  or  a  circular  caisson  in  a  given 
shaft.  The  circular  shape  is  easier  to  build  and  sink,  and, 
owing  to  arch  action,  thinner  walls  can  be  used.  No  horizon- 
tal reenforcement  or  cross-braces  are  needed,  and  therefore 


FIG.  16.  —  Shoe,  Lower  Part  of  Form,  and  Reinforcement,  Rondout  Caisson 

a  grab  bucket  can  be  used  to  advantage.  The  rectangu- 
lar shape,  on  the  other  hand,  requires  less  excavation,  and 
the  walls  of  any  caisson  must  be  made  thicker  than  are 
needed  for  strength  to  give  weight  for  sinking.  In  general 
it  is  probable  that  the  circular  shape  is  better  for  a  one-  or 
two-compartment  shaft,  and  for  a  three-compartment  shaft 
in  very  bad  ground;  the  rectangular  for  a  three-compart- 
ment shaft  in  ordinary  soft  ground.  For  long  shafts,  a 
rectangular  caisson  is  a  necessity. 

An  allowance  should  always  be  made  for  a  possible 
tilt  from  the  vertical  that  may  amount  to  from  18  in.  to  2  ft., 


SINKING  THROUGH  SURFACE  49 

either  by  battering  or  stepping  the  walls  on  the  inside,  or 
by  making  the  caisson  larger  than  the  neat  size  required. 
In  a  rectangular  caisson  the  side  walls  should  be  braced  with 
temporary  struts  while  sinking,  and  the  permanent  buntons 
or  cross-walls  placed  after  it  has  reached  its  final  bearing 
on  the  rock;  it  may  otherwise  be  impossible  to  line  the  guides 
up  plumb. 

The  relative  costs  of  piling,  forepoling,  and  caisson  are 
influenced  by  local  conditions  and  by  the  type  of  shaft 
desired.  For  instance,  a  permanent  shaft,  such  as  the 
D.,  L.  &  W.  shaft  at  Wilkesbarre,  must  have  a  masonry 


FIG.  17.  — -  Derrick  and  Head  Works  Rondout  Caisson 

lining  through  the  surface  anyhow,  and  it  is  therefore  not 
fair  to  charge  against  the  excavation  the  cost  of  the  concrete 
in  the  caisson.  In  a  temporary  shaft,  on  the  other  hand, 
the  excess  of  the  cost  of  the  caisson  over  the  cost  of  a  timber 
lining  must  be  charged  against  the  excavation.  The  writer 
believes,  nevertheless,  that  wherever  the  ground  is  not 
firm  enough  to  support  itself  for  one  set  in  advance  of  the 
timber  lining,  a  caisson  is  safest  and  cheapest  in  the  long 
run.  A  possible  exception  may  be  made  to  this  statement 
in  the  case  of  a  moderate  depth  of  very  wet  ground  where  the 
work  is  done  by  a  contractor  who  owns  the  equipment  for 
driving 'steel  piles  and  can  recover  them  after  the  masonry 
lining  is  completed  and  use  them  on  other  work. 


50  PRACTICAL  SHAFT  SINKING 

The  costs  given  below  should  be  fairly  representative 
for  the  different  methods  of  work : 

1.  Shaft  excavated  14  X  20£  ft.  through  6  ft.  of  soil 
and  14  ft.  of  quicksand,  not  very  wet.  Sides  supported  by 
2-in.  oak  sheeting  driven  by  mauls  and  braced  by  five  sets 
of  10  X  12  in.  timber. 


FIG.  18.  —  Dump,  Rondout  Caisson,  Showing  Flowing  Nature  of  Material 

Per  Foot      Per  Cubic  Yard 

Labor $27.25  $2.57 

Lumber,  6600  feet  B.  M.  at  $30 9.90  .93 

Erection  of  derrick,  etc 3.00  .29 

Superintendence 3.00  .29 

Sundry 2.00  .18 

Coal  and  pumping  5.00  .47 

Total $50.15  $4.73 

2.  Shaft  excavated  12  X  20  ft.  3  in.  through  45  ft,  of 
clay  and  gravel.  Sides  supported  by  sets  of  10  X  10  in. 
pine  timber  spaced  4f-ft.  centers  and  hung  from  top.  1^-in. 
lagging: 

Per  Foot      Per  Cubic  Yard 

Labor $19.50  $2.17 

Lumber,  240  feet  per  foot  at  $25 6.00  .66 

Bolts,  15  pounds  per  foot  at  $0.03 45  .05 

Erection  of  head-frame,  etc 2.00  .22 

Superintendence 2.00  .22 

Power 1.50  .17 

Sundry 1.00  .11 

Total    .                                                     ..  $32.45  $3.60 


SINKING  THROUGH  SURFACE  51 

3.  Shaft  excavated  15  X  37  ft.  through  21  ft.  of  dry 
sand.     Sides  supported  by  interlocking  steel  sheet  piling 
driven  with  steam  hammer  and  braced  with  sets  of  8  x  10  in. 
timber : 

Labor  Costs  Only  Per  Foot  Per  Cubic  Yard 

Driving  sheeting $  6.55  $  .32 

Removing  sheeting 1.85  .09 

Timbering 2.05  .10 

Excavation    8.20  .40 

Total $18.65  $  .91 

The  cost  of  superintendence,  sundries,  and  plant  rental 
would  amount  to  about  $10  per  ft.  or  .50  per  yard  at  a  low 
estimate,  and  the  cost  of  the  steel  sheet  piling,  if  charged 
entirely  to  this  job,  would  amount  to  $110  per  ft.,  or  $5.30 
per  yard. 

4.  Caisson  26  ft.  outside  diameter,  21  ft.  inside  diameter, 
sunk  through  56  ft.  semi-liquid  mud  and  boulders: 

Per  Cubic  Yard 
Per  Foot  Excavation 


[  Materials   

$  27.00 

$1.35 

Concrete  •{  Labor    . 

7.00 

.35 

[  Forms  and  shoe  

,*'     23.00 

1.15 

Sinking  caisson  

;       38.00 

1.90 

Plant  erection  .",..•  

3.00 

.15 

Superintendence  

..*...         5.00 

.25 

Sundry  

......         5.00 

.25 

Coal  and  power 

6.00 

.30 

Total    

$114.00 

$5.70 

SEALING    THE    CAISSON  TO   ROCK 

After  the  cutting  edge  of  a  caisson  has  reached  rock,  it 
is  still  necessary  to  construct  a  seal  to  permanently  exclude 
sand  and  water.  Often  a  stratum  of  stiff  clay  or  disin- 
tegrated shale  is  found  under  the  soft  material  and  imme- 
diately over  the  rock.  If  this  occurs  the  cutting  edge  will 
sink  into  it,  automatically  shutting  out  water  until  a 
concrete  wall  is  built;  if  not,  the  making  of  the  seal  will  be 
very  troublesome. 


52 


PRACTICAL    SHAFT    SINKING 


Running  mud  may  be  checked  long  enough  to  allow  it 
to  harden  by  caulking  under  the  shoe  with  blocks  of  wood 
and  old  sacks.  Streams  of  water  and  quicksand  require  a 
wedging  curb  of  some  kind.  The  English  method  of  sealing 
tubbing  to  rock  can  be  applied;  this  may  be  done,  Fig.  19, 
by  cutting  out  the  rock  under  the  shoe  until  the  cutting 
edge  attains  a  fair  bearing  all  around,  then  driving  numerous 
wooden  wedges  into  the  crack  until  the  wrater  is  blocked 
back.  Another  plan,  Fig.  20,  is  to  lead  the  water  to  the 
center  of  the  shaft  through  pipes  set  opposite  the  main 


#•;&•£ 


FIG.  19 


FIG.  20 


feeders.  A  brick  or  concrete  wall  is  then  built  from  the 
rock  to  the  caisson,  surrounding  the  pipes  and  forcing  all 
the  water  through  them.  When  this  masonry  has  set  hard 
enough  to  stand  the  water  pressure,  the  pipes  are  plugged. 
Several  small  pipes  should  also  be  built  into  the  wall  so  that 
grout  can  be  pumped  back  of  it  to  take  up  small  leaks. 
With  either  method  great  care  is  necessary,  in  commencing 
the  rock  excavation,  to  avoid  opening  up  a  new  leak. 

Sometimes  the  quantity  of  water  may  be  so  great  that 
it  cannot  be  shut  off  as  described.  If  it  is  anticipated  that 
the  ground  will  be  very  wet,  provision  should  be  made  in 
the  design  of  the  caisson  for  the  use  of  compressed  air 
as  described  below.  This  provision  was  made  in  the 


SINKING  THROUGH   SURFACE 


f>3 


FIG.  21.  —  Steel  Shoe  for  D.,  L.  &  W.  Caisson 


FIG.  22.  —  Shoe  and  Form  for  Bottom  of  D.,  L.  &  W.  Caisson 


54  PRACTICAL  SHAFT  SINKING 

D.,  L.  &  W.  caisson  referred  to  above,  although  air  was 
not  used. 

The  following  notes  on  D.,  L.  &  W.  caisson,  taken  from  the 
Engineering  News  for  September  28,  1908,  are  of  interest. 

In  sinking  the  shaft,  after  the  surface  had  been  removed 
with  plows  and  scrapers  and  the  bottom  of  the  excavation 


FIG.  23.  —  Showing  Reenforcement  for  D.,  L.  &  W.  Concrete  Caisson 

made  perfectly  level,  a  steel  shoe,  shown  in  Fig.  21,  was 
placed  on  the  bottom  of  the  excavation.  This  was  made  of 
f-in.  plate,  was  24  in.  wide,  32  in.  high,  and  reenforced,  as 
shown,  with  riveted  angles.  The  shelf  which  formed  the 
base  for  the  concrete  was  placed  8  in.  above  the  toe  of  the 
vertical  plate.  The  outside  dimensions  of  the  cutting  shoe 
were  28  X  59  ft.  5  in.  The  outside  form  for  the  concrete 
was  built  up  flush  with  the  outside  edge  of  the  shoe.  The 


SINKING  THROUGH  SURFACE 


55 


inside  form  at  the  bottom  was  inclined  as  shown  in  Fig.  22, 
being  given  a  batter  until  the  wall  was  7  ft.  thick  on  the 
sides  and  5  ft.  on  the  ends,  when  vertical  forms  were  put  in 
place.  The  concrete  was  reenforced  with  tie-rods,  as  shown 
in  Fig.  23,  and  the  walls  were  decreased  in  thickness  in 
steps,  as  shown  in  Fig.  25,  until  they  reached  a  uniform 
thickness  of  2  ft.  8  in.  at  the  top.  When  a  height  of  20  ft. 
of  the  concrete  was  reached,  the  bottom  forms  were  removed 


FIG.  24.  —  D.,  L.  &  W.  Caisson  Ready  for  Sinking 

and  the  concrete  caisson  then  carefully  leveled  preparatory 
to  sinking.  In  order  to  provide  for  the  contingency  of 
having  to  resort  to  compressed  air  in  sinking  in  case  the 
inflow  of  water  proved  too  great  to  be  handled  by  pumps, 
arrangements  were  made  to  put  in  an  air  deck  in  case  of 
necessity.  Sinking  was  carried  on  day  and  night,  and  the 
excavating  gang  consisted  of  a  foreman  and  sixteen  men 
to  each  shift.  The  materials  were  hoisted  in  buckets  by 
means  of  derricks,  as  shown  in  Fig.  24.  Just  as  the  caisson 
reached  the  rock  which  was  being  cleaned  off  preparatory 
to  putting  in  the  seal,  the  river  rose  and  the  shaft  was 


PRACTICAL  SHAFT  SINKING 


Longitudinal  Section. 

FIG.  25.  —  Plan  and  Sections  of  D. 


Crofts     Section. 

L.  &  W.  Caisson  Walls 


SINKING  THROUGH  SURFACE 


57 


flooded.  It  was  found  impossible  to  pump  it  out  and  the 
shaft  was  allowed  to  fill,  to  remain  full  until  the  river  had 
subsided.  When  the  caisson  had  sunk  to  the  level  of  the 
rock,  it  was  found  that  a  temporary  seal  would  have  to  be 
put  in  place  during  the  construction  of  the  permanent  seal. 
This  temporary  seal  was  made  of  yellow  pine  blocks, 
12  x  12  in.  in  size,  and  wedges  a,  Fig.  26.  Six-inch  bleeder 
pipes  were  left  to  drain  off  the  water  while  the  seal  was  being 


FIG.  26 

put  in  place.  The  pipes  were  closed  after  the  temporary 
seal  had  been  completed. 

To  provide  a  place  for  the  permanent  seal  it  was  necessary 
to  take  out  rock  for  a  depth  of  20  ft.,  so  as  to  build  a  wall 
to  carry  the  caisson.  During  the  blasting  of  this  rock  great 
care  had  to  be  taken  to  prevent  jarring  out  the  temporary 
seal.  As  the  rock  was  being  excavated  a  grout  was  forced 
back  of  the  temporary  blocking  by  means  of  a  grout  pump 
with  an  air  pressure  of  80  Ibs. 

In  order  to  give  a  firm  footing  for  the  concrete  wall  the 


58  PRACTICAL  SHAFT  SINKING 

rock  was  recessed  so  as  to  form  a  toe  for  the  wall,  and  in 
order  to  give  good  contact  between  the  underpinning  wall 
and  the  caisson,  the  lower  edge  of  the  caisson  was  roughened. 
Fig.  26  shows  the  method  of  making  the  permanent  seal 
between  the  caisson  and  the  concrete  foundation.  The 
concrete  c  was  put  in  place  as  soon  as  possible  after  the  rock 
excavation  had  been  completed,  and  was  of  the  form  shown. 
Next  the  ring  of  concrete  d  was  placed,  and  grout  was  then 
pumped  into  the  pipes  after  the  concrete  c  and  d  had  set. 
The  final  wedge  of  concrete  e  was  laid  after  the  concrete 
lower  down  had  set  and  everything  below  had  been  made 
thoroughly  tight,  the  edge  between  e  and  d  being  caulked 
after  the  pipe  had  been  grouted;  g  is  broken  stone  packed 
in  between  a  brick  dam  and  the  wooden  seal;  the  brick  dam 
is  intended  to  lead  the  water  to  the  pipes  6. 


CHAPTER  IV 

SINKING  THROUGH  SOFT  GROUND — PNEUMATIC  PROCESS  — 
SHIELD  METHOD. 

SOFT  GROUND 

FOR  sinking  through  soft  ground  containing  more  water 
than  can  be  pumped,  the  three  methods  referred  to  in 
Chapter  III  have  been  developed  in  this  country  and  abroad. 
They  may  be  described  as  follows: 

THE    PNEUMATIC    PROCESS 

The  pneumatic  caisson  is  an  application  of  the  principle 
of  the  diving  bell  that  has  been  widely  used  for  founding 
deep  piers.  It  is  also  used  for  soft-ground  shafts,  particu- 
larly construction  shafts  from  which  tunnels  are  to  be  driven 
under  compressed  air.  A  caisson  is  constructed  similar 
to  the  open  caissons  already  described,  except  that  an  air- 
tight deck  is  built  over  the  entire  opening  8  or  10  ft.  above 
the  cutting  edge,  Fig.  26.  The  deck  is  made  strong  enough 
to  resist  an  air  pressure  equivalent  to  the  hydrostatic  head 
at  tne  depth  which  the  caisson  is  expected  to  reach.  One 
or  more  openings  in  the  deck  are  provided,  fitted  with  air 
locks  which  retain  air  pressure  but  permit  the  entrance  of 
men  and  the  removal  of  spoil. 

The  caisson  is  constructed  above  the  surface  and  sunk 
by  excavating  under  the  cutting  edge  as  in  the  open  type. 
The  air  pressure  is  raised  as  the  caisson  sinks  and  is  always 
kept  slightly  in  excess  of  the  water  pressure  at  the  cutting 
edge.  Water  is  absolutely  excluded  —  no  matter  how  wet 
and  soft  the  material  the  work  is  done  in  the  dry.  In  this 
way  shafts  can  be  sunk  through  river  silt  and  flowing  quick- 
sands that  cannot  be  handled  in  the  open. 

59 


60  PRACTICAL  SHAFT  SINKING 

The  cost  of  excavation  under  compressed  air  is  in 
general  much  higher  than  that  of  open  work.  In  the  first 
place  grab  buckets  cannot  be  hoisted  through  an  air  lock, 
so  hand  digging  is  necessary;  second,  a  special  class  of  high- 
priced  laborers  must  be  employed  whose  wages  increase 
with  the  depth,  while  the  length  of  the  shifts  must  be 
reduced;  third,  the  air  locks  applicable  to  caissons  are  costly 
to  build,  and  as  their  construction  is  covered  by  patents 
controlled  by  one  or  two  corporations,  they  are  quite  costly 
to  rent;  fourth,  the  masonry  of  the  caissons  must  be  made 
very  heavy  to  overcome  the  upward  pressure  of  the  air. 
Some  grounds,  however,  can  be  " blown"  out  of  the  caisson 
and  very  little  digging  is  necessary;  in  this  case  excavation 
is  cheap.  The  limit  for  pneumatic  sinking  in  loose  ground 
is,  in  general,  100  ft.  below  water  level,  as  men  cannot 
stand  a  pressure  greater  than  that  corresponding  to  this 
depth. 

The  deck  in  a  shaft  caisson  must  be  removable,  and  is, 
therefore,  made  of  timber  or  steel,  fitted  into  a  recess  left 
in  the  wall.  Two  openings  should  be  provided,  one  in  the 
middle  for  the  excavation^  lock  and  another  for  the  man 
lock.  American  locks  are  now  standardized  to  fit  a  36-in. 
circular  opening.  Small  openings  must  also  be  made  in 
the  deck  for  the  air  connections  and  the  "blow  pipe." 

A  man  lock  consists  of  a  steel  cylinder,  about  4  ft.  in 
diameter  and  8  ft.  long,  with  flanged  head.  Eighteen-inch 
openings  in  the  head  are  fitted  with  doors  which  swing 
downwards  in  opening,  and  close  against  a  rubber  gasket. 
A  small  hole  in  each  head,  closed  by  a  stop-cock  inside  the 
lock,  permits  the  entrance  of  compressed  air  from  the 
caisson  and  its  escape  to  the  atmosphere.  The  lower  door 
and  stop-cock  being  closed,  the  upper  door  is  opened  and 
several  men  enter  the  lock.  The  upper  door  is  then  closed 
from  outside,  and  a  lock-tender  standing  inside  the  lock 
closes  the  upper  and  opens  the  lower  stop-cock.  When 
the  pressure  in  the  lock  has  become  equal  to  that  in  the 
caisson,  he  opens  the  lower  door  and  the  men  climb  down  a 


SINKING  THROUGH  SOFT   GROUND 


61 


ladder  to  the  bottom.     In  letting  men  out  the  process  is 
reversed. 

The  locking  through  of  men  is  the  most  precarious  part 
of  compressed-air  work.  Too  quick  an  application  of  pres- 
sure causes  "blocking  of  the  ears"  —  intolerable  pain  in 
the  ears  and  head  due  to  unequal  pressure  on  the  two  sides 
of  the  ear  drum  —  and  too  quick  a  reduction  of  pressure 


Hoist  Rope.- 
Stuffing  Box. 


2  Air  Outlet.       Doors. 


FIG.  27.  —  Pneumatic  Caisson 

may  cause  the  "bends"  —  caisson  workers'  paralysis  — 
always  dangerous  and  sometimes  fatal.  For  pressures 
below  20  Ibs.,  men  accustomed  to  the  work  can  be  locked 
through  safely  in  two  or  three  minutes  and  can  work  eight- 
hour  shifts;  at  higher  pressures  the  locking  time  must  be 
increased  and  the  length  of  the  shifts  decreased.  With 
45  Ibs.  of  air,  forty-five-minute  shifts  are  worked  and 
twenty-five  minutes  must  be  taken  in  locking  through. 


62  PRACTICAL  SHAFT  SINKING 

The  principle  of  the  excavation  lock  is  the  same  as  that 
of  the  man  lock,  but  the  doors  and  valves  are  all  controlled 
from  outside.  The  patented  " straight-through"  lock  is 
the  best  type,  in  fact  the  only  satisfactory  type  for  caisson 
work.  Several  forms  are  made,  all  of  which  require  the 
bucket  to  be  hoisted  on  a  single  rope.  In  the  ''pot-lid" 
lock  the  rope  passes  through  a  stuffing-box  in  the  middle  of 
the  upper  door,  which  is  literally  a  lid  fastened  to  the  lock 
by  six  heavy  bolts  hinged  to  the  lock  and  engaging  slots 
in  the  edge  of  the  door.  The  door  is  carried  by  a  buffer 
at  the  lower  end  of  the  rope.  When  a  bucket  is  lowered 
into  the  lock,  the  upper  door  is  also  lowered  on  to  its  seat. 
The  lock-tender,  who  stands  outside,  raises  the  hinge  bolts 
into  their  slots  and  tightens  the  nuts  before  opening  the  lower 
door.  In  consequence  of  the  continual  tightening  and 
loosening  of  the  bolts  this  lock  is  rather  slow  in  operation, 
but  it  is  very  simple. 

Other  forms  of  the  straight-through  lock  have  the  upper 
door  in  two  halves  which  close  upon  a  stuffing-box  on  the 
rope.  In  this  way  only  the  stuffing-box  is  lifted  instead 
of  the  entire  door,  and  the  operation  is  much  quicker. 
In  one  type  the  doors  are  operated  by  high-pressure  air 
cylinders. 

The  air  pumped  into  the  caisson  by  the  compressors 
escapes  by  forcing  its  way  out  through  the  ground  close 
under  the  cutting  edge.  As  it  is  often  necessary  to  excavate 
several  feet  below  the  cutting  edge  to  sink  the  caisson,  some 
provision  must  be  made  to  remove  the  water  that  collects 
in  the  depression.  This  can  of  course  be  done  with  a  pump 
driven  by  high-pressure  air,  but  it  is  also  possible  to  blow 
the  water  out  directly.  A  4  to  6  in.  pipe,  closed  by  a  stop- 
cock at  the  lower  end,  is  led  through  the  deck  and  out  over 
the  top  of  the  caisson,  and  a  suction  hose  reaching  the 
sump  is  attached  to  the  lower  side  of  the  stop-cock.  An 
opening,  closed  by  a  small  valve  or  a  wooden  plug,  is  made 
in  the  pipe  above  the  stop-cock.  When  the  stop-cock  is 
open  the  air  pressure  lifts  the  water  into  the  pipe;  the  small 


SINKING  THROUGH  SOFT  GROUND  63 

valve  is  opened  at  the  same  time  and  a  quantity  of  air  flows 
in  and  mixes  with  the  stream  of  water,  decreasing  its  specific 
gravity  until  the  weight  of  the  whole  column  of  water  is 
less  than  the  air  pressure.  The  water  is  then  driven  com- 
pletely out  of  the  caisson.  By  replacing  the  valve  with  a 
small  high-pressure  air  connection  it  has  been  possible  to 
raise  water  out  of  a  caisson  70  ft.  deep  with  13  Ibs.  of  air. 

Fine  sand  and  silt  containing  much  water  can  be  blown 
out  through  the  pipe,  and  caissons  have  been  sunk  without 
hoisting  a  bucket  of  dirt. 

The  use  of  compressed  air  makes  it  very  much  easier 
to  seal  the  caisson  to  rock.  There  is  of  course  no  trouble 
about  keeping  the  water  out ;  the  difficulty  is  to  prevent  the 
air  from  blowing  the  grout  out  of  the  concrete,  leaving  it 
porous.  One  plan  is  to  lay  a  strip  of  heavy  duck  over  the 
crack,  nailing  one  edge  to  the  caisson  and  the  other  to  the 
rock.  Concrete  is  then  laid  on  top  of  this  duck. 

Some  kinds  of  very  wet  ground  possess  considerable 
viscosity.  In  these  the  pneumatic  process  can  be  worked 
to  a  greater  depth  than  is  theoretically  possible  by  reducing 
the  pressure  and  blowing  out  the  mud  and  water  that  flow 
in  under  the  cutting  edge.  One  example  of  this  has  been 
cited  —  where  water  was  raised  70  ft.  with  13  Ibs.  of  air; 
in  England  recently  several  piers  were  founded  at  a  depth 
of  130  ft.  below  water  level  with  45  Ibs.  of  air  pressure. 
Under  such  conditions  the  difference  between  the  hydrostatic 
pressure  and  the  air  pressure  is  accounted  for  by  internal 
friction  of  the  water  and  the  ground.  It  is  probable  also 
that  the  actual  hydrostatic  head  is  Deduced  by  the  ah* 
bubbles  which  escape  under  the  cutting  edge  into  the  sur- 
rounding ground. 

THE    SHIELD   METHOD 

Shields,  similar  in  principle  to  those  so  extensively  used 
for  subaqueous  soft-ground  tunnels,  have  also  been  applied 
to  soft-ground  shafts,  Fig.  28.  A  shoe  is  constructed  with 
a  cutting  edge  slightly  larger  than  the  outside  of  the  com- 


64 


PRACTICAL  SHAFT  SINKING 


pleted  shaft  lining;  a  vertical  lap  plate  or  shield  is  attached 
to  the  outer  perimeter  of  the  shoe,  and  a  number  of  screw 
(or  hydraulic)  jacks  are  set  on  top  of  the  shoe  and  inside 
the  shield  plate.  The  frame  of  the  shoe  is  sometimes  made 
of  wood,  but  steel  is  preferable.  The  shield  is  made  of 
i  to  f-in.  plate  iron  and  extends  from  18  in.  to  3  ft.  above 


!   Cross  Bunfons 


FIG.  28.*  —  Shield  Method  of  Sinking 

the  top  of  the  shoe  proper.     The  method  of  operation  is  as 
follows:  • 

As  soon  as  the  shaft  is  started,  bearing  timbers  or  trusses 
are  constructed  to  hang  the  lining  from  as  previously 
described  in  connection  with  forepoling.  The  shoe  is 
assembled  in  place  with  jacks  screwed  down,  and  the  shaft 
lining  is  completed  from  the  surface  to  the  heads  of  the 

*  Figs.  28,  34,  36,  37,  and  47  are  reproduced  from  the  copyrighted  instruc- 
tion papers  and  bound  volumes  of  the  International  Correspondence  Schools 
by  special  permission  of  the  International  Textbook  Company. 


SINKING  THROUGH  SOFT  GROUND  65 

jacks.  Excavation  is  then  started  and  the  shoe  sinks  until 
enough  distance  is  gained  to  allow  another  course  of  lining 
to  be  placed  beneath  the  completed  section  and  inside 
the  shield.  If  the  jacks  are  used  to  force  the  shoe  down, 
they  must  be  withdrawn  before  the  course  of  lining  can  be 
placed. 

The  upper  edge  of  the  shield  must  always  be  kept  above 
the  lower  edge  of  the  completed  lining,  and  to  insure  this  in 
bad  ground  it  is  necessary  to  hang  the  shoe  from  the  bearing 
timbers  with  chains  and  ratchet  jacks.  Sometimes  shoes 
are  made  so  that  the  opening  can  be  completely  closed 
with  steel  plates  to  prevent  an  inrush  of  sand. 

Tunnels  driven  with  shields  are  circular  and  lined  with 
rings  of  cast-iron  segments  2  ft.  wide.  Many  European 
shafts  are  lined  this  way,  but  the  American  shafts  to  which 
the  shield  method  has  been  applied  are  rectangular  and 
lined  with  "skin-to-skin"  timbers  or  plank  laid  flat. 

The  chief  disadvantage  of  a  shield,  even  at  a  moderate 
depth,  is  its  liability  to  hang  up  on  a  boulder  on  one  side 
while  the  other  side  settles,  thus  wedging  itself  and  throwing 
the  shaft  out  of  line.  This  tendency  can  be  largely  overcome 
by  the  proper  suspension  of  the  shield,  but  the  depth  which 
can  be  reached  is  limited  when  the  ground  is  soft  and  wet 
enough  to  exert  fluid  pressure.  At  100  ft.  below  ground- 
water  level,  for  example,  the  pressure  of  wet  quicksand  will 
at  least  be  45  Ibs.  per  square  inch,  sufficient  to  force  enough 
sand  and  water  to  flood  the  shaft  through  a  very  small 
opening.  It  is  impossible  to  jack  a  closed  shoe  down, 
displacing  the  ground  under  it. 


CHAPTER  V 

SINKING  IN  ROCK  —  ARRANGEMENT  OF  HOLES  —  TOOLS  AND 
METHODS  USED  IN  DRILLING  —  COSTS  AND  SPEED 

DYNAMITE  and  the  power  drill  have  made  solid  rock  the 
easiest  material  through  which  to  sink  a  shaft,  and  prac- 
tically all  American  mining  shafts  are  in  rock  for  the  greater 
part  of  their  depth.  As  has  been  said  before,  hand  sinking 
is  the  cheapest  and  quickest  method;  although  a  boring 
process  has  been  developed,  it  is  only  applied  where  such 
immense  quantities  of  water  are  encountered  that  hand 
sinking  is  impossible. 

Outside  of  the  boring  process,  the  improvements  in  rock 
sinking  have  all  related  to  breaking  the  rock  and  hoisting  it. 
No  practicable  mechanical  excavator  or  loader  has  yet  been 
devised.  Grab  buckets  that  work  well  in  soft  ground  are 
failures  in  blasted  rock.  A  steam  shovel,  useful  in  a  tunnel, 
is  of  course  out  of  the  question  in  the  bottom  of  a  sinking 
shaft. 

Drilling  and  Blasting.  —  The  universal  method  of  shaft 
sinking  in  rock  is  to  drill  a  number  of  holes  in  the  bottom, 
charge  them  with  dynamite  and  shoot  them,  and  to  load 
the  broken  rock  by  hand  into  shaft  buckets  which  are  then 
hoisted  out.  When  all  the  loose  rock  has  been  removed 
the  process  is  repeated.  As  it  is  very  difficult  to  drill  holes 
through  loose  rock,  the  broken  material  must  be  all  removed 
before  the  next  round  of  holes  is  started.  This  creates  an 
additional  difficulty  for  the  mechanical  digger,  for  while  a 
grab  might  be  made  to  remove  most  of  the  loose  rock  after 
a  blast,  hand  work  would  still  have  to  be  resorted  to  to  get 
the  bottom  ready  for  drilling. 

Shafts  are  drilled  on  the  ''center-cut"  principle.    Eight 


SINKING   IN   ROCK 


67 


or  ten  holes  are  drilled  on  a  slant,  separated  at  the  top  but 
converging,  thus  forming  a  wedge  known  as  the  "sump." 
"Reliever,"  or  bench,  holes  are  drilled  back  of  the  sump 
holes,  each  row  being  more  nearly  vertical;  the  end  or  out- 
side holes  point  slightly  away  from  the  vertical  and  toward 
the  wall  line  of  the  shaft.  The  sump  is  first  shot  and  the 
broken  rock  removed  or  "mucked"  out,  forming  a  cavity 
into  which  the  bench  rounds  can  be  successively  shot. 
All  muck  should  be  removed  before  each  succeeding  round 
is  shot. 


FIG.  29.  —  Shaft  (a) 


Two  systems  of  drilling  and  mucking  exist.  In  the  first 
the  holes  for  the  entire  cut  —  sump  and  benches  —  are 
drilled  at  one  time,  the  sump  is  shot,  and  then  the  benches 
as  required.  In  the  second  the  sump  only  is  drilled  and 
shot,  and  the  benches  are  drilled  while  the  sump  is  being 
mucked.  The  first  plan  is  particularly  applicable  to  small 
shafts  and  to  circular  shafts;  a  rectangular  or  elliptical 
shape  is  needed  to  give  room  for  simultaneous  drilling  and 
mucking. 

Fumeless,  or  gelatine,  dynamite  should  in  all  cases  be 
used  for  underground  work.  The  fumes  from  ordinary 
glycerine  dynamite  make  it  imposssible  for  the  men  to  get 
back  to  work  promptly  after  a  shot.  The  strength  of  the 
dynamite  used  depends  on  the  character  of  the  rock,  but 


PRACTICAL  SHAFT  SINKING 


40  and  60  per  cent,  gelatine  are  the  most  common  strengths 
used. 

The  number  and  depth  of  the  holes  and  the  quantities 
of  powder  loaded  vary  so  greatly  with  the  size  of  the  shaft 
and  the  nature  of  the  rock  that  no  general  rules  can  be  stated. 
The  systems  actually  used  at  several  shafts  were  as  follows: 

(a)  Shaft  13  X  26  ft.,  through  Western  Pennsylvania 
coal  measures:  Shale,  slate,  and  limestone;  horizontal 
stratification;  holes  as  in  Fig.  29;  40  per  cent,  gelatine: 


Number 

Depth 

Inclination 
with 
Vertical 

Loaded 
with 

Sump 

8 

Feet 

10 

Degrees 
35 

Pounds 
4 

Relievers 

8 

8 

25 

3 

Benches  
End  
Total  charge  

8 
8 

8 
8 

0 
10  back 

2^ 
2£ 
96 

Average  gain  per  cut,  6  feet. 

Average  gain  per  week  of  19  shifts,  24  feet  (no  timber). 

Mucking  and  drilling  simultaneous;  2  drills  used  on  1  bar. 

(6)  Shaft  14  X  48  ft.,  through  anthracite  measures:  Red 
sandstone;  stratification  horizontal;  holes  as  in  Fig.  30; 
40  per  cent,  gelatine: 


Number 

Depth 

Inclination 

Loaded 
with 

Sump    
Relievers    
Benches    
End  '  
Total  charge  per  round    

8 
8 

24 

8 

Feet 
10 
8 
8 
8 

Degrees 
35 
25 
lOtoO 
10  back 

Pounds 

5 

4 
3 
3 
168 

Average  gain  per  cut,  6  feet. 

Average  gain  per  week  of  18  shifts,  16  feet. 

Mucking  and  drilling  simultaneous;  2  drills  used  on  1  bar. 

(c)  Shaft  10  X  22  ft.,  through  quartz  conglomerate 
(Shawangunk  grit) ;  horizontal  stratification,  but  very  few 
bedding  planes;  holes  as  in  Fig.  31;  60  per  cent,  gelatine: 


SINKING   IN   ROCK 


Number 

Depth 

Inclination 

Loaded 
with 

Sump  

8 

Feet 
10 

Degre 
35 

Pounds 
3-i 

Sump 

4 

8 

o 

Ql 

Relievers  
Benches  
End  ,  
Total  charge  per  round 

8 
8 
8 

9 

8 
8 

25 
0 
10  back 

21 
2 
2 
94 

Average  gain  per  cut,  5£  feet. 

Average  gain  per  week  of  20  shifts,  22  feet. 

Mucking  and  drilling  simultaneous;  5  drills  used  on  2  bars. 

The  four  additional  sump  holes  shown  were  used  on 
account  of  extra  hardness  of  the  rock. 

(d)  Shaft  elliptical,  19  ft.  4  in.  X  33  ft.,  through  West 
Virginia  coal  measures:  Hard  gray  sandstone;  40  per  cent, 
gelatine;  holes  as  in  Fig.  32;  horizontal  stratification: 


Number 

Depth 

Inclination 

Loaded 
with 

Feet 

Degrees 

Pounds 

Sump 

10 

12 

35 

5 

Relievers    

8 

10 

25 

4 

Benches   

14 

10 

10 

4 

End  

6 

10 

10  back 

3 

Total  charge  per  round  

— 

— 

— 

156 

Average  gain  per  cut,  8  feet. 

Average  gain  per  week  of  20  shifts,  18  feet. 

Mucking  and  drilling  simultaneous;  3  drills  used  on  1  long  bar,  1  short  bar. 

(e)  Shaft  circular,  17  ft.  diameter,  through  Hamilton 
and  Marcellus  shales:  Rock  distorted;  stratification  irregular; 
holes  as  in  Fig.  33, but  about  45  degrees;  60  per  cent,  gelatine: 


Number 

Depth 

Inclination 

Loaded 
with 

Feet 

Degrees 

Pounds 

Sump           

6 

8 

35 

2| 

8 

6 

20 

u 

Rib  

16 

6 

10  back 

1 

Total  charge  per  round  

— 

— 

— 

43 

Average  gain  per  cut,  5£  feet. 

Average  gain  per  week  of  19  shifts,  33  feet. 

All  drilling  on  one  shift,  mucking  on  two  shifts;  5  drills  used  on  5  tripods. 


70 


PRACTICAL  SHAFT  SINKING 


Drilling  Tools.  —  Hand  drilling,  once  universal,  has  been 
entirely  superseded  in  the  United  State  by  compressed-air 
drilling,  and  it  is  in  fact  difficult  to  obtain  hammer  men. 


FIG.  31.  —  Shaft  (c) 

In  other  countries  where  labor  is  cheap,  drilling  is  still  done 
by  hand.     In  the  commonest  method,  a  drill  or  " jumper" 


FIG.  32.  —  Shaft  (d) 


FIG.  33.  —  Shaft  (e) 


of  1-in.  steel  is  turned  by  one  man  and  struck  by  one 
or  two  others  with  8-lb.,  double-faced  hammers,  Fig.  34a. 
Americans  and  Europeans  use  a  30-in.  stiff  handle;  the 
Southern  negro  prefers  to  "drive  steel"  with  a  slightly  longer 


SINKING   IN   ROCK 


71 


handle  whittled  down  until  it  bends  like  a  whip.  Jumpers 
are  given  a  single  cutting  edge,  usually  curved,  Fig.  40. 
Two  men  should  strike  each  steel  wherever  practicable,  as 
they  can  obviously  drill  twice  as  fast  as  a  single  striker  at 
three-fourths  the  cost.  As  much  depends  on  the  man  that 
turns  the  steel  as  on  the  striker,  for  considerable  skill  is 
needed  to  produce  a  round,  straight  hole.  Three  good 
men  can  drill  If -in.  holes  in  hard  sandstone  at  the  rate  of 
2  ft.  per  hour. 


FIG.  34a 


FIG.  34 

In  the  Tyrol  a  system  of  single-handed  drilling  has  been 
developed.  The  driller  turns  a  light  steel  with  one  hand 
and  wields  a  4-lb.  hammer,  Fig.  346,  with  the  other.  A  skil- 
ful man  can  thus  drill  3-ft.  holes  quite  rapidly,  but  the  holes 
are  too  small  for  regular  shaft  sinking. 

For  slate,  churn  drills,  Fig.  38,  are  often  used.  The  drill 
consists  of  a  straight  bar,  6  to  12  ft.  long,  with  a  bit  at  each 
end.  An  iron  weight  is  sometimes  welded  around  or  forged 
into  the  drill  2  ft.  from  one  end,  thus  increasing  the  weight 


72 


PRACTICAL  SHAFT  SINKING 


of  the  drill  without  increasing  its  length  or  diameter.  The 
drill  is  handled  by  two  or  three  men.  When  the  weighted 
drill  is  used,  the  hole  is  started  with  the  short  end,  and  when 
it  has  reached  a  depth  of  2  ft.  the  drill  is  reversed. 


FIG.  35 

The  reciprocating,  compressed  air  drill  is  the  most  widely 
used  machine  for  drilling  rock.  It  was  first  put  into  prac- 
tical use  by  Mr.  Fowle,  of  Boston,  in  the  construction  of 
the  Hoosac  tunnel,  and  since  then  has  steadily  grown  in 
popularity.  It  is  turned  out  by  the  thousands  by  the 


SINKING  IN  ROCK  73 

Ingersoll-Rand  Co.,  the  Sullivan  Machinery  Co.,  the 
McKiernan  Drill  Co.,  and  others,  and  although  each  maker 
has  certain  features  of  his  own,  especially  in  the  valve 
arrangement,  the  general  design  is  standardized  and  the 
general  features  are  shown  in  Fig.  36.  Piston  and  rod  are 
turned  out  of  a  single  billet  of  special  steel,  and  to  the  end 
of  the  rod  the  drill  steel  is  rigidly  attached  by  a  U-bolt 
chuck.  The  cylinder  is  made  of  cast-iron  and  slides  longi- 
tudinally in  a  guide  frame  (or  shell)  clamped  to  the  drill 
mounting.  As  the  drill  cuts  into  the  rock,  the  cylinder  is 
fed  forward  by  a  square-thread  screw  mounted  on  the 
frame.  The  piston  is  rotated  mechanically  by  a  "rifle  bar" 
and  ratchet  so  that  the  cutting  edge  of  the  bit  will  not  strike 
two  successive  blows  in  the  same  spot.  This  rotative  effect 
is  necessary  to  drill  a  round  hole. 

The  machine  commonly  used  for  shaft  sinking  has  a 
3|-in.  cylinder  and  a  6|-in.  stroke,  weighs  280  Ibs.,  and  will 
drill  down  holes  in  hard  rock  at  the  rate  of  about  7  ft.  per 
hour,  including  time  lost  in  changing  steels.  The  length  of 
feed  is  24  in.,  hence  the  drills  must  be  changed  every  2  ft. 
The  starter  is  2  ft.  long  beyond  the  shank  (the  portion  of  the 
drill  grasped  by  the  chuck),  and  the  following  steels  are  4  ft., 
6  ft.,  8  ft.,  etc.,  respectively.  (See  Fig.  39.)  Drill  steels 
are  usually  sharpened  with  a  +  bit,  although  X  bits  and 
straight  I  bits  are  sometimes  used.  Where  a  large  number 
of  drills  are  in  operation  a  sharpening  machine  may  be  used 
to  advantage. 

Two  types  of  valve  motion  can  be  obtained.  In  the 
first,  the  valve  which  controls  the  piston  is  thrown  by  a 
tappet  struck  by  the  piston  itself;  the  Rand  " Little  Giant" 
is  an  example  of  this.  In  the  second,  a  piston  valve  is  used 
which  is  thrown  by  a  difference  in  air  pressure  on  the  two 
ends.  The  Sullivan  "Slugger"  is  a  drill  of  this  type.  The 
Sergeant  drill  is  a  compromise,  having  an  auxiliary  valve, 
driven  by  contact  with  the  piston,  which  governs  the  air 
pressure  on  the  ends  of  the  main  piston  valve.  The  Slugger 
type  strikes  a  hard,  uncushioned  blow  and  is  adapted  to 


74 


PRACTICAL  SHAFT  SINK  INC, 


SINKING   IN   ROCK  75 

use  in  hard  rock  with  compressed  air.  Wet  steam  will  not 
operate  the  valve  readily,  and  the  drill  is  slow  if  wet  steam 
is  used.  The  tappet  drill,  on  the  other  hand,  having  a  posi- 
tive valve  motion  will  do  good  work  on  wet  steam.  Its 
blow  is  slightly  cushioned.  The  auxiliary-valve  drill  strikes 
a  hard  blow,  will  run  by  steam,  and  in  addition  has  the 
advantage  of  a  variable  stroke.  This  feature  makes  it 
easier  to  start  a  hole. 

A  good  many  types  of  air-hammer  drill  have  been 
recently  developed,  and  have  replaced  the  reciprocating 
types  for  light  work.  In  these  the  drill  steel  is  struck  by  a 
reciprocating  hammer  and  has  very  little  motion  of  its 
own.  The  drill  steel  is  hollow,  and  the  powdered  rock  is 
blown  out  of  the  hole  by  a  portion  of  the  exhaust  air  led 
to  the  cutting  face  through  the  hole  in  the  steel.  The 
Water  Leyner  drill,  Fig.  35,  which  is  now  built  to  compete 
with  the  larger  sizes  of  reciprocating  drills,  works  on  the 
hammer  principle.  In  this  the  cuttings  are  removed  by  a 
stream  of  water  pumped  through  the  hollow  steel  to  the 
cutting  face;  a  portion  of  the  exhaust  air  is  allowed  to  mix 
with  the  water.  This  drill  has  made  some  remarkable 
records  in  hard  rock  tunnels  in  the  West.  A  great  advan- 
tage of  the  drill  is  that  no  dust  is  created  in  drilling  up-holes 
in  tunnels,  making  this  work  very  much  more  healthy  for 
the  drill  runners. 

In  rectangular  shafts,  drills  are  mounted  on  "  shaft  bars," 
or  single  screw  columns,  Fig.  37.  The  drill  itself  is  held  by 
a  clamp,  which,  when  its  bolts  are  loosened,  can  be  slid 
along  or  revolved  around  the  bar,  at  the  same  time  per- 
mitting the  drill  to  be  swung  sidewise  to  any  angle.  When 
the  clamp  bolts  are  tightened  the  drill  is  rigidly  held  in  posi- 
tion. The  bar  is  set  horizontally  across  the  shaft,  wooden 
blocking  being  used  to  form  a  good  bearing  between  its  ends 
and  the  walls  of  the  shaft.  Two  and  sometimes  three  drills 
are  mounted  on  each  bar.  "Column  arms,"  used  for  off- 
setting the  drill,  do  not  work  satisfactorily  with  a  shaft  bar, 
and  besides  are  unnecessary  in  a  rectangular  shaft. 


76 


PRACTICAL  SHAFT  SINKING 


In  circular  shafts  it  is  difficult  to  cover  the  area  to  be 
drilled  with  a  straight  bar,  and  in  the  writer's  opinion  it  is 
best  to  mount  the  drills  on  tripods,  Fig.  34.  The  tripod, 
while  it  possesses  all  the  adjustability  of  other  forms  of 
mounting,  is  less  rigid  and  more  cumbersome.  To  do  good 
work  all  loose  rock  should  be  removed  and  the  legs  set  on 
solid  rock.  This  feature,  however,  is  not  objectionable  in 
a  shaft  where  mucking  and  drilling  are  not  carried  on  simul- 
taneously. 


Before  blasting,  the  drills  and  mountings  must  of  course 
be  hoisted  out  of  the  shaft.  In  England,  a  drilling  frame 
for  use  in  circular  shafts  has  been  patented.  This  consists 
of  a  ring  from  which  six  bars,  on  which  the  drills  are  mounted, 
project  radially.  The  ring  is  supported  by  legs  and  held 
rigid  by  jack-screws  in  the  ends  of  the  six  bars.  The  chief 
advantage  of  this  frame  is  that  it  is  unnecessary  to  detach 
all  the  drills  before  blasting;  the  whole  can  be  hoisted  off 
the  bottom  and  hung  in  the  shaft  without  hindering  the 


SINKING   IN  ROCK 


77 


78    *  PRACTICAL  SHAFT  SINKING 

passage  of  the  bucket  which  passes  up  and  down  through  the 
ring.  Another  advantage  of  the  frame  is  that  the  manifold 
is  attached  to  it,  the  drills  being  connected  to  the  manifold 
by  short  pieces  of  hose.  It  is  thus  necessary  to  have  only 
one  main  air  hose  hanging  in  the  shaft  instead  of  a  small 
hose  for  each  machine. 

The  best  grade  of  canvas  and  rubber  hose  should  be 
used  for  the  drills,  wire  wound  for  air,  and  marlin  wound 
for  steam.  " Steam  hose"  should  be  ordered  always  for 
use  with  steam,  as  "air  hose"  will  not  stand  heat.  A  sec- 
tion of  1-in.  hose  50  ft.  long  is  used  for  operating  each  drill. 

Operation.  —  Shaft  sinking  is  usually  carried  on  twenty- 
four  hours  a  day.  The  inside  work  is  done  by  three  shifts 
of  men  working  eight  hours  each,  the  outside  by  three 
8-hour  or  two  12-hour  shifts.  The  12-hour  outside  shift 
is  customary  in  the  coal  fields;  elsewhere,  the  8-hour  shift 
for  every  one  is  prevalent.  Shifts  are  usually  changed  at 
7  A.M.  and  3  and  11  P.M.,  sometimes  an  hour  later.  The 
men  are  given  twenty  minutes  for  lunch  in  the  middle  of 
each  shift. 

Wages  vary  with  the  locality,  but  in  general  men  are 
paid  better  for  drilling  and  mucking  in  a  shaft  than  in  any 
other  kind  of  rock  excavation.  On  account  of  the  high 
wages  paid  in  America  machine  drilling  is  universal,  and  the 
shifts  are  limited  to  the  number  of  men  that  can  be  worked 
to  the  best  advantage.  Speed  is  not  attempted  at  the 
expense  of  efficiency.  In  South  Africa,  on  the  other  hand, 
Kaffir  labor  is  cheap,  hand  drilling  is  usual,  and  as  many 
men  are  worked  as  the  shafts  will  hold. 

The  great  depth  of  the  shafts  on  the  Rand  makes  the 
highest  possible  speed  desirable,  even  at  an  increased  cost. 
In  both  countries  speed  is  increased  without  an  increase  of 
cost  by  the  payment  of  a  bonus  to  the  sinkers  as  a  reward 
for  additional  progress. 

The  size  of  the  shifts  for  any  given  shaft  depends  upon 
the  number  of  drills  required  and  upon  the  experience  and 
ability  of  the  sinkers  obtainable.  With  first-class  men, 


SINKING  IN  ROCK  79 

the  men  on  each  shift  at  the  13  X  26  ft.  shaft  referred  to 
before  as  (a)  would  be  as  follows: 

Inside  men,  8  hours:  One  shift  boss,  at  $3;  two  drillers, 
at  $2.75;  two  helpers,  at  $2.50;  six  muckers,  at  $2.25. 

Outside  men,  12  hours:  One  engineer;  one  head  tender; 
three  car  men  on  dump;  one  fireman;  one  compressor 
man. 

General  outside,  10  hours:  One  foreman;  one  mechanic; 
two  carpenters  (on  timber) ;  one  blacksmith  and  helper. 

The  17-ft.  circular  shaft  (e)  would  require: 

Drilling  shift:  One  shift  boss;  five  drillers;  five  helpers; 
one  extra  man. 

Mucking  shifts:  One  shift  boss;  nine  muckers. 

Outside  same  as  shaft  (a). 

In  South  African  shafts,  which  are  usually  about  9  X  26 
ft.,  when  drilling  is  done  by  hand,  each  shift  consists  of  one 
white  shift  boss  and  about  35  Kaffir  laborers  who  drill  or 
muck  as  may  be  required. 

Thorough  organization  is  essential  to  progress  and 
economy.  Each  man  must  know  his  place  and  take  it 
without  losing  time  in  getting  started.  Any  condition 
that  prevents  systematic  work  is  fatal  to  economy.  For 
instance  an  inflow  of  water,  sufficient  to  cause  a  loss  of  time 
after  every  blast  while  the  bottom  is  being  pumped  dry, 
will  lessen  the  rate  of  sinking  far  more  than  can  be  calculated 
by  adding  together  the  actual  delays. 

Ventilation.  —  Foul  air  and  powder  smoke  in  the  shaft 
bottom  hinder  work  almost  as  much  as  water.  As  a  rule 
vertical  shafts  ventilate  themselves  surprisingly  well  to  a 
depth  of  400  to  500  ft.,  but  at  greater  depths  and  sometimes 
at  much  lesser  depths,  artificial  ventilation  must  be  resorted 
to.  The  cheapest  method,  where  the  natural  draft  needs 
only  a  slight  assistance,  is  an  "air  box,"  or  wooden  pipe 
carried  up  one  compartment  of  the  shaft;  into  this  box  is 
turned  a  jet  of  air  or  steam  or  the  exhaust  of  the  pump,  if 
one  is  used.  The  box  is  built  of  1  X  12  in.  boards.  Another 
way  to  help  natural  ventilation  in  a  rectangular  shaft  is  to 


80  PRACTICAL  SHAFT  SINKING 

divide  the  shaft  into  two  compartments  with  a  brattice 
attached  to  a  row  of  bun  tons;  one  compartment  will  then 
establish  itself  as  an  upcast,  the  other  as  a  downcast.  If 
all  steam  pipes  to  pumps  are  kept  in  one  compartment  this 
action  is  certain  to  occur.  In  any  case  steam  pipes  should 
be  kept  together  in  one  end  of  the  shaft. 

In  deep  shafts  positive  ventilation  is  best  assured  by  the 
use  of  a  fan  or  blower  discharging  into  a  large  air  pipe 
carried  down  the  shaft  and  lengthened  from  time  to  time 
as  the  shaft  deepens.  A  15-in.  standard  volume  blower, 
engine-  or  motor-driven,  is  sufficient  to  ventilate  a  shaft  to 
any  ordinary  depth.  The  pipe  may  be  made  of  boards, 
canvas,  or  light,  galvanized  sheet  iron.  The  latter,  although 
more  expensive  than  wood  or  canvas,  is  air-tight  and  is  not 
liable  to  injury  from  concussion. 

Progress.  —  Progress  in  shaft  sinking  is  influenced  by  so 
many  different  conditions  —  quality  of  rock,  size  and  shape 
of  the  shaft,  presence  or  absence  of  water,  efficiency  of  labor 
and  plant  —  that  •  it  is  very  hard  to  make  any  general 
statements  concerning  it.  The  best  progress  records  are 
made  in  deep  rectangular  shafts  on  the  Rand,  in  Transvaal, 
South  Africa.  These  shafts,  as  has  been  said  before,  have 
a  section  about  9  X  26  ft.  hi  the  rock;  work  is  carried  on 
by  three  8-hour  shifts  7  days  a  week;  two  compartments  are 
used  for  hoisting,  and  every  man  that  can  be  worked  is  put 
into  the  shaft.  Kaffir  labor  is  not  only  cheap,  but  the 
Kaffir  will  work  under  conditions  of  crowding  to  which  a 
white  man  will  not  submit.  The  records  made  in  several 
South  African  shafts  are  given  in  Table  1 :  the  average 
progress  is  seen  to  be  about  135  ft.  per  month,  and  the 
maximum  213.5  ft. 

The  progress  in  this  country  under  normal  conditions 
ranges  from  60  to  80  ft.  for  rectangular  timbered  shafts, 
although  very  much  higher  speeds  are  sometimes  reported. 
The  soft  shale  in  the  coal  fields  of  the  Middle  Western  states 
is  easy  to  drill  and  shoot.  Good  records  are  made  in  Kansas 
and  Southern  Illinois.  An  account  of  a  shaft  near  Atchison, 


SINKING   IN   ROCK  81 

Kan.,  published  in  the  Engineering  and  Mining  Journal 
for  July  26,  1902,  states  that  the  daily  progress  in  soft  shale 
was  7  ft.  No  monthly  figures  were  given.  A  good  record 
was  recently  made  on  a  17-ft.  circular  shaft  in  the  "  Hudson 
River  shale"  (dark  blue  sandstone  and  sandy  slate)  on  the 
New  York  Aqueduct.  The  system  of  drilling  and  muck- 
ing used  at  this  shaft  is  described  above  under  (e) ;  .the  rock 
was  quite  hard  but  broke  readily.  The  average  progress 
made  here  is  shown  in  the  tabulation  of  American  shafts, 
Table  2.  The  best  month's  work  is  177  ft.  No  work  was 
done  on  Sundays.  The  writer  believes  this  to  be  a  record 
for  American  shafts.* 

The  rate  of  progress  in  the  European  circular  shafts  lies 
between  the  American  and  the  African  rate.  The  rock 
penetrated  is  in  general  softer  than  that  found  in  this 
country. 

Tables  1  and  2  give  the  dimensions  of  a  number  of  shafts 
and  the  progress  made  in  them.  Wherever  obtainable,  the 
nature  of  the  rock  penetrated  and  the  cost  per  fo'ot  is  given. 
The  figures  were  obtained  from  various  articles  in  the 
technical  papers,  from  the  proceedings  of  various  mining 
institutes,  and  from  the  writer's  own  records;  some  of  the 
South  African  data  were  taken  from  the  "Deep  Level  Mines 
of  the  Rand,"  by  G.  A.  Denny,  1902. 

Cost.  —  Cost  figures  cover  a  wider  range  than  progress 
figures  and  are  harder  to  get.  The  cheapest  shaft  on  record 
is  the  one  near  Atchison,  referred  to  above,  the  cost  of  which, 
as  stated,  was  $7  per  foot.  This  cost  stands  alone  in  its 
glory  as  the  tabulated  figures  show.  Mr.  Henry  Rawie 
published  in  Mines  and  Minerals  an  itemized  statement  of 
the.  costs  of  a  shaft  sunk  in  West  Virginia,  in  1906.  These 
ran  as  follows: 

*  Since  the  above  was  written,  the  Breakneck  Shaft. on  the  New  York 
Aqueduct  was  sunk  183  ft.  in  a  month.  The  rock  was  hard  granite.  The 
system  used  was  the  same  as  at  No.  1  Moodna,  but  six  3f  drills  on  tripods 
were  used  on  the  drilling  shift.  One  mucking  shift  only  was  worked  on 
Sunday;  no  drilling  shift. 


82  PRACTICAL  SHAFT  SINKING 

HOIST  SHAFT,  14  FT.  X  22  FT.,  180  FT.  DEEP 

Per  Foot 

Labor,  sinking  and  timbering    $24.70 

Plant    5.55 

Superintendence 

Explosives 3.88 

Coal 2.55 

Timber 6.67 

Miscellaneous 5.55 

$48.90 

The  sinking  costs  of  a  pair  of  shafts  sunk  in  Western 
Pennsylvania  a  year  later  were  as  follows  : 

HOIST  SHAFT,  13  FT.  X  26  FT.,  422  FT.  DEEP 

Per  Foot 

Labor,  sinking $51.00 

Plant    2.40 

Superintendence 4.35 

Explosives  2.75 

Coal   5.50 

Oil 60 

Freight 50 

Miscellaneous 7.90 

Total -.... $75.00 

AIR  SHAFT,  13  FT.  X  22  FT.,  383  FT.  DEEP 

Per  Foot 

Labor,  sinking $57.50 

Plant    2.40 

Superintendence 4.90 

Explosives 3.00 

Coal  6.05 

Oil 60 

Freight 50 

Miscellaneous 7.14 

Total    $82.09 

Water  per  minute:  Hoist  shaft,  50  gallons;  air  shaft, 
120  gallons. 

Costs  have  risen  greatly  in  the  last  decade,  since  no  sub- 
stantial improvements  in  methods  or  machinery  have  been 
made  to  offset  the  increase  in  wages.  Contract  prices  are 


SINKING   IN   ROCK 


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84  PRACTICAL  SHAFT  SINKING 

not  generally  obtainable,  as  most  shafts  are  put  down  by 
private  corporations,  but  prices  high  enough  to  include  a 
good  profit  to  the  contractor  eight  to  ten  years  ago  would 
not  cover  his  costs  to-day. 

Twenty-five  shafts  ranging  in  depth  from  350  to  over 
1000  ft.  are  required  for  the  portion  of  the  New  York  Aque- 
duct now  under  construction.  All  but  two  of  these  have 
been  let  by  contract,  and  the  bid  prices  are  public  prop- 
erty. The  bids,  however,  are  unbalanced  in  every  case,  and 
do  not  give  a  fair  idea  of  shaft  prices.  They  range  from 
$175  to  $350  per  foot. 


CHAPTER  VI 

THE  SINKING-DRUM  PROCESS.       MAMMOTH  PUMP.       THE 
FREEZING  PROCESS 

Sinking-drum  Process.  —  For  sinking  through  the  very 
great  depths  of  water-bearing  sand  and  clay  that  exist  in 
some  of  the  German  mining  districts,  a  method  has  been 
developed  that  does  not  require  any  hand  work  in  the  shaft 
bottom  until  the  lining  is  completed  to  rock.  The  shafts 
are  necessarily  circular  and  are  lined  with  cast-iron  tubbing. 
A  heavy  masonry  caisson,  with  an  inside  diameter,  somewhat 
greater  than  that  of  the  finished  shaft  desired,  is  first  con- 
structed and  sunk  for  50  or  60  ft.,  as  described  in  a  previous 
chapter.  If  the  ground  beneath  the  cutting  edge  is  sufficiently 
firm,  it  is  leveled  off  and  a  foundation  ring  of  masonry  built 
under  the  tapered  part  of  the  wall.  If  not,  a  concrete  floor 
is  laid  over  the  bottom  (under  water  if  necessary)  and  the 
caisson  is  pumped  out.  A  heavy  iron  ring  projecting  inside 
the  face  of  the  wall  all  around  is  built  into  the  masonry 
foundation  ring  or  into  a  groove  cut  in  the  wall  above  the 
concrete  floor.  A  second  ring,  placed  near  the  top  of  the 
caisson,  is  connected  to  the  first  with  heavy  iron  rods,  and 
the  space  between  the  rings  and  around  the  rods  is  filled 
with  masonry,  forming  an  inner  tube.  The  upper  ring  pro- 
jects inside  this  inner  tube  and  serves  as  a  base  for  a  circle 
of  powerful  hydraulic  jacks  acting  downward. 

A  very  strong  cast-steel  shoe  or  cutting  edge  with  an 
outside  diameter  slightly  less  than  the  inside  diameter  of  the 
tube  is  then  assembled  on  the  shaft  bottom,  and  rings  of 
cast-iron  tubbing  bolted  together  are  built  up  from  the  top 
of  the  shoe  to  the  heads  of  the  jacks.  If  a  concrete  floor  has 
been  laid  it  is  broken  up  with  a  huge  churn  drill,  excavation 

85 


86  PRACTICAL  SHAFT  SINKING 

is  started  with  a  grab  bucket  or  some  other  mechanical 
digger,  and  the  shoe  and  tubbing  commence  to  sink  of  their 
own  weight.  The  inner  masonry  lining  acts  as  a  guide  for 
the  iron  sinking  drum,  and  must  therefore  be  built  with  its 
axis  exactly  vertical,  correcting  any  deviation  that  may  have 
occurred  in  sinking  the  caisson.  The  rubbing  surface  is 
usually  formed  by  I-beams  built  into  the  masonry. 

When  motion  ceases,  the  jacks  are  brought  into  play  and 
the  drum  is  forced  down,  additional  rings  of  tubbing  being 
built  up  under  the  heads  of  the  jacks.  When  the  first  drum 
can  be  forced  no  farther,  the  bottom  of  the  shaft  is  plugged 
with  concrete,  the  water  is  pumped  out,  and  a  second  sinking 
drum  built  up  inside  the  first.  The  concrete  is  broken  up  as 
before,  the  jacks  shifted  so  as  to  engage  the  top  of  the  inner 
drum,  and  sinking  is  resumed.  As  many  as  four  drums 
have  been  used,  reaching  in  one  place  a  depth  of  508  ft. 

The  three  methods  used  for  excavation  under  water 
inside  the  drum  are: 

The  Grab  Bucket.  —  The  action  of  a  clam-shell  or 
orange-peel  bucket  is  too  well  known  to  require  explanation. 
Either  bucket  will  handle  coarse  sand,  gravel,  and  boulders 
to  advantage,  but  will  not  retain  fine  wet  sand  through  a 
long  hoist,  and  will  not  dig  tough  clay. 

The  Sack  Borer.  —  This  is  a  gigantic  auger  with  the  rigid 
stem  extending  up  the  center  of  the  shaft,  Fig.  41.  The 
stem  is  constructed  of  heavy  flanged  pipes  bolted  together, 
and  is  terminated  at  the  upper  end  by  a  splined  section 
which  serves  as  the  shaft  of  a  large  horizontal  worm-wheel. 
A  hoisting  rope,  leading  to  a  powerful  engine  and  attached 
to  a  swivel  at  the  top  of  the  splined  section,  suspends  the 
borer. 

The  stem  is  turned  by  an  engine  acting  through  the 
worm-gear  and  its  worm,  and  is  lowered  gradually  by  the 
hoisting  rope.  When  the  top  of  the  splined  section  reaches 
the  worm-gear,  it  is  disconnected  from  the  stem  proper  and 
raised,  and  another  section  of  standard  pipe  is  added  beneath 
it.  Cross-arms,  fitted  with  rollers  at  their  ends,  are  attached 


THE   SINKING-DRUM   PROCESS 


87 


to  the  stem  at  intervals;  the  rollers  bear  against  the  sides  of 
the  completed  shaft  and  prevent  the  stem  from  buckling. 
The  material  cut  by  the  borer  is  collected  in  two  heavy 


FIG.  41.  —  Sack  Borer 

canvas  sacks  fastened  to  the  backs  of  the  cutters.  Formerly 
they  were  rigidly  attached,  and  the  whole  apparatus  was 
hoisted  every  time  the  sacks  were  filled.  Now,  however, 
the  sacks  are  mounted  on  frames  sliding  on  two  pairs  of 


88  PRACTICAL  SHAFT  SINKING 

guides  attached  to  the  cross-arms  on  the  stem,  and  are 
hoisted  by  light,  independent  engines.  The  sack  borer 
is  adapted  to  clay  and  sand. 

The  Mammoth  Pump.  —  This  is  an  application  of  the 
air  lift,  used  in  conjunction  with  a  percussion  borer  or  large 
churn  drill,  Fig.  43.  A  discharge  pipe  A,  open  at  both  ends, 
is  carried  down  along  the  boring  rod  from  the  surface  and 
is  terminated  just  above  the  point  of  the  borer.  A  com- 
pressed-air pipe  B  is  also  carried  down  the  rod  and  connected 
into  the  discharge  or  suction  pipe  A  near  the  bottom.  The 


FIQ.  42.  —  Construction  of  Sinking  Drum  for  Hydraulic  Flushing  Process 

borer  being  in  operation,  the  air  is  turned  on  and  a  stream 
of  water,  mud,  and  sand  is  lifted  through  the  discharge  pipe. 
The  pump  will  handle  practically  any  material  that  will 
enter  the  discharge  pipe. 

The  chief  difficulty  with  the  sinking  drum  has  been  the 
thickness  of  iron  required  to  withstand  the  earth  pressure 
at  great  depth,  and  uncertain  strains  caused  by  boulders 
under  the  cutting  edge.  The  internal  flanges  on  the  tubbing 
cannot  be  made  very  wide  without  interfering  with  the  free 
passage  of  boring  tools  in  the  shaft,  hence  the  strength  of 
the  lining  depends  on  its  thickness  alone.  This  has  reached 
3£  in.,  and  at  that  thickness  collapse  has  occurred  in  several 


THE   MAMMOTH  PUMP 


FIG.  43.  —  Compound  Drum  and  Mammoth  Pump  and  Borer 
A,  Suction  Pipe  and  Overflow  of  Muddy  Water;  B,  Compressed  Air; 
C,  Masonry  Caisson;  D,  Shoe,  Acting  on  Anchoring  Ring;  E,  Anchor  Rods. 


90  PRACTICAL  SHAFT  SINKING 

cases.  Further  increase  is  not  practicable  on  account  of 
the  weight  of  the  segments  and  the  difficulty  of  handling 
them.  The  cost  would  also  be  excessive. 

The  compound  sinking  drum  (patented  in  Germany  by 
Mr.  Pattberg)  is  a  decided  improvement.  In  this,  occa- 
sional rings  of  tubbing  are  provided  with  broad  internal 
flanges,  and  the  space  between  these  is  filled  with  concrete 
or  brick,  leaving  the  interior  of  the  shaft  perfectly  smooth. 
The  masonry  not  only  strengthens  the  tubbing,  but  also 
adds  weight  where  it  will  do  the  most  good,  and  expedites 
sinking. 

The  friction  and  adhesion  between  the  ground  and  the 
drum  have  been  lessened  by  hydraulic  flushing.  For  this, 
the  shoe  and  three  or  four  rings  of  tubbing  immediately 
above  it  are  made  slightly  larger  than  the  rest  of  the  lining. 
In  the  upper  side  of  the  shoulder  thus  formed,  water  passages 
are  provided  which  are  connected  to  a  pressure  pump. 
While  sinking,  the  pump  is  operated  and  the  drum  is  par- 
tially surrounded  by  a  film  of  water.  This  expedient  has 
been  very  successful.  (See  Fig.  42.) 

The  sinking  drum  is  sealed  to  the  solid  measures  by  forc- 
ing the  cutting  edge  into  them  by  the  full  power  of  the  jacks. 
If  necessary  the  shaft  can  be  bored  into  the  rock  by  the 
Kind-Chaudron  method,  as  will  be  explained  later.  The 
entire  process  will  probably  be  made  clearer  by  a  short 
description  of  an  actual  piece  of  work. 

Shaft  5,  of  the  Rheinpreussen  Colliery,  Homburg-am- 
Rhein,  Germany,  was  expected  to  penetrate  nearly  500  ft. 
of  quicksand  and  mud.  Sinking  was  started  with  a  brick 
caisson  C,  Fig.  43,  29.2  ft.  inside  diameter,  with  walls  about 
3.5  ft.  thick.  This  reached  a  depth  of  65  ft.  Nine  feet  of 
concrete  was  placed  on  the  bottom  under  water  and  the  shaft 
pumped  out.  The  anchor  ring  D,  anchor  rods  E,  and  pres- 
sure ring,  designed  for  a  maximum  pressure  of  3000  tons, 
were  erected  and  a  brick  inner  lining  built  around  the  rods, 
reducing  the  diameter  of  the  shaft  to  25.68  ft. 

A  compound  drum  F,  with  an  outside  diameterof  25.52 


THE  FREEZING   PROCESS  91 

ft.,  and  a  diameter  inside  the  broad  flanges  and  the  brick 
lining  of  21.32  ft.,  was  now  built  and  sinking  was  started  with 
a  percussion  borer  and  mammoth  pump.  The  concrete  was 
bored  through  in  four  days,  and  an  average  advance  of 
about  5  ft.  a  day  was  made  in  the  soft  ground.  The  progress 
was,  in  fact,  limited  by  the  rate  at  which  tubbing  and  walling 
could  be  built  up  under  the  heads  of  the  jacks. 

It  was  possible  to  force  the  compound  drum  to  a  depth  of 
245  ft.  The  shaft  was  then  filled  for  60  ft.  with  sand  and 
gravel  instead  of  concrete,  was  pumped  out,  and  an  inner 
iron  drum,  3^  in.  thick  and  19.35  ft.  in  inside  diameter,  was 
built  up  to  the  jacks.  This  drum  stuck  at  a  depth  of  315  ft., 
60  ft.  of  gravel  was  again  filled  into  the  shaft,  and  a  third 
drum,  17.38  ft.  in  inside  diameter,  was  built  up  to  the  jacks. 
This  was  forced  to  a  depth  of  343  ft.,  where  the  cutting  edge 
stuck  in  clay  solid  enough  to  permit  the  shaft  to  be  pumped 
out.  A  fourth  drum,  15.3  ft.  in  inside  diameter,  was  then 
built,  which  reached  the  solid  coal  measures  at  a  depth  of 
508  ft. 

Shaft  4  was  sunk  simultaneously,  with  exactly  similar 
drums.  The  third  drum  reached  a  depth  of  433  ft.  before 
the  shaft  could  be  pumped  out.  The  completion  of  both 
shafts  to  the  rock  took  three  years. 

The  Freezing  Process.  —  The  great  depth  to  which  frost 
penetrates  the  ground  in  Siberia  and  other  cold  countries 
enables  shafts  to  be  sunk  through  soft  ground  to  consider- 
able depths  during  the  winter  months.  Continued  freezing 
gives  the  sides  all  the  support  that  is  necessary  until  rock 
is  reached  and  a  permanent  lining  built  up. 

It  occurred  to  F.  H.  Poetsch  in  1883  that  this  condi- 
tion could  be  imitated  artificially.  His  method  is  to  bore 
a  number  of  holes  around  and  somewhat  outside  of  the 
periphery  of  the  proposed  shaft,  and  to  case  them  through 
the  soft  strata  to  the  rock.  A  freezing  plant  is  erected  at  the 
shaft  head,  and  the  brine  or  freezing  solution  is  circulated 
down  interior  pipes  and  up  through  the  bore-hole  casings 
until  the  surrounding  ground  is  frozen  to  a  solid  mass.  The 


92  PRACTICAL  SHAFT  SINKING 

holes  are  bored 'about  3  ft.  apart;  the  form  of  the  frozen 
ground  is  consequently  cylindrical.  At  first  the  cylinder 
is  hollow,  but  as  the  freezing  continues,  it  gradually  becomes 
solid  ice.  Excavation  is  then  commenced,  the  frozen  ma- 
terial being  loosened  with  picks  or  light  charges  of  explo- 
sives. 

In  Europe  70  or  80  shafts  have  been  sunk  by  the  freezing 
process,  the  thickness  of  the  soft  ground  in  some  cases 
reaching  300  ft.  Most  of  these  shafts  are  in  France,  Bel- 
gium, or  Germany;  a  few  have  been  frozen  in  England  by 
continental  contractors.  In  the  United  States  the  process 
has  so  far  found  a  very  limited  application.  One  shaft  in 
Michigan  was  frozen  through  about  100  ft.  of  quicksand, 
and  an  unsuccessful  attempt  was  made  to  freeze  a  shaft  in 
Pennsylvania. 

A  number  of  European  shafts  started  by  the  freezing 
method  have  been  completely  lost  through  some  accident. 
Notwithstanding  this,  the  method  is  being  improved  and 
greater  and  greater  depths  are  attempted  and  reached. 
Water-bearing  rock  strata  are  successfully  frozen.  A  shaft 
in  Belgium  has  been  sunk  by  freezing  through  700  ft.  of 
soft  ground  and  wet  rock. 

A  detailed  description  of  the  freezing  process,  written  by 
Mr.  Sidney  F.  Walker,  may  be  found  in  the  August,  1909 
issue  of  Mines  and  Minerals. 

The  chief  difficulties  met  with  in  freezing,  especially  in 
deep  freezing,  are  deviation  of  the  bore  holes,  salts  in  solu- 
tion in  the  ground  water,  bursting  freezing  pipes,  and  the 
tendency  of  ice  to  flow  under  pressure.  The  first  trouble 
can  be  met  by  measuring  the  drift  of  the  holes,  and  by  boring 
additional  holes  when  the  divergence  of  those  already  bored 
is  too  great.  Salt  solutions  are  of  course  very  hard  to  freeze, 
and  their  presence  in  the  ground  necessitates  a  much  longer 
freezing  period  than  would  otherwise  be  necessary.  A  burst 
pipe  allows  the  freezing  solution  itself  to  flow  into  the  ground, 
forming  a  soft  spot  that  it  is  almost  impossible  to  freeze  at 
all.  The  obvious  way  to  prevent  this  is  to  use  very  strong 


THE  FREEZING   PROCESS  93 

tested  pipe,  and  it  is  now  found  advisable  net  to  circulate 
the  freezing  solution  through  the  bore  hole  casing  itself,  but 
to  insert  an  inner  and  outer  freezing  tube  and  to  withdraw 
the  casing.  The  flowage  of  ice  cannot  be  prevented  and 
limits  the  depth  for  which  the  freezing  process  is  feasible. 
Hard  freezing  checks  this  tendency. 

A  freezing  period  long  enough  to  thoroughly  solidify  the 
ground  is  the  first  essential  for  successful  sinking.  The 
smallest  crack  or  seam  which  will  admit  a  few  drops  of  water 
will  soon  enlarge  itself  until  a  disastrous  break-through 
occurs.  It  is  also  necessary,  from  time  to  time  as  the  shaft 
is  excavated,  to  support  the  sides  with  some  form  of  sus- 
pended lining. 

The  Anhalt  Government  Salt  Mine,  at  Leopolds-Hall, 
Stassfurt,  is  an  example  of  a  successful  application  of  the 
freezing  process.  Drilling  was  commenced  early  in  1899 
and  26  holes  about  5  in.  in  outside  diameter  were  drilled  on  a 
26.25  ft.  circle  and  cased  to  a  depth  of  325  ft.  The  drilling 
was  difficult,  and  was  not  completed  until  June,  1900.  By 
this  time  the  freezing  plant  (which  consists  at  most  shafts 
of  two  75  horse-power  ammonia  compressors)  was  ready 
and  it  was  started  June  22.  Sinking  was  commenced  on 
September  2,  and  on  September  19,  at  a  depth  of  30  ft.,  a 
small  leak  which  existed  in  the  middle  of  the  bottom  broke 
through  and  flooded  the  shaft.  Freezing  was  then  continued 
until  the  end  of  November,  and  sinking  was  again  started. 
By  the  end  of  February,  1901,  202  ft.  had  been  sunk,  and 
the  shaft  was  then  lined  with  iron  tubbing.  The  space 
between  the  tubbing  and  the  rock  was  filled  with  concrete 
mixed  with  a  solution  of  calcined  soda.  Periods  of  sinking 
and  lining  then  alternated,  until  on  July  4,  the  shaft  was 
lined  complete  to  the  bottom  of  the  frozen  wall.  The  total 
time  required  for  a  depth  of  325  ft.  was  therefore  two  and 
one-half  years,  an  average  progress  of  11  ft.  per  month. 

The  Chapin  Mine  Co.,  Iron  Mountain,  Mich.,  decided  to 
sink  a  shaft  in  the  center  of  a  small  valley  crossing  its  prop- 
erty. Attempts  to  sink  by  ordinary  methods  having  failed, 


94  PRACTICAL  SHAFT  SINKING 

in  1887  a  contract  was  let  to  the  Poetsch-Sooy smith  Freezing 
Co.  to  sink  the  shaft  by  the  freezing  process.  At  the  site  of 
the  shaft  the  rock  was  covered  by  95  ft.  of  quicksand, 
gravel,  and  boulders.  The  sand  had  some  clay  mixed  with 
it,  contained  1  per  cent,  of  water,  and  would  flow  almost  like 
water. 

The  installation  of  the  freezing  pipes  was  performed  by 
the  Chapin  Mine  Co.  itself,  and  was  finished  in  the  summer 
of  1888.  Twenty-six  10-in.  bore  holes,  spaced  evenly  on 
29-ft.  circle,  were  driven  and  cased  to  rock,  great  difficulty 
being  experienced  in  keeping  them  vertical  on  account  of 
the  boulders.  Eight-inch  freezing  pipes  f  in.  thick,  flush 
inside  and  out  and  closed  at  the  bottom,  were  lowered  into 
the  holes  and  the  casings  were  then  withdrawn.  Inner 
tubes  1£  in.  in  diameter  were  lowered  into  the  freezing 
tubes,  their  lower  ends  being  kept  8  in.  above  the  bottom. 
The  upper  ends  of  the  tubes  were  connected,  as  shown  in 
Fig.  44,  to  the  brine  pipes  of  a  50-ton-per-day  capacity 
Linde  freezing  plant  operated  by  a  55  horse-power  engine, 
driven  by  compressed  air.  Two  hundred  cubic  feet  of  brine 
consisting  of  a  25  per  cent,  solution  of  calcium  chloride  was 
used  for  the  freezing  fluid,  the  entire  quantity  making  a 
circuit  every  thirty-three  minutes. 

Excavation  and  timbering  were  started  fifteen  days  after 
freezing  was  begun  and  were  continued  for  seventy-eight 
days,  when  rock  was  reached  in  one  end  of  the  shaft.  During 
this  period  two  and  a  half  days  were  lost  by  the  interrup- 
tion of  the  air  supply  to  the  engine.  Some  water  had  been 
finding  its  way  up  through  the  unfrozen  core  in  the  middle 
of  the  shaft;  the  quantity  now  increased  and  sand  began  to 
come  in  with  it.  The  shaft  was  at  once  filled  with  water 
and  an  additional  freezing  pipe  put  down.  Four  months 
and  a  half  after  freezing  was  begun  the  shaft  was  sunk 
7  or  8  ft.  into  the  rock.  At  this  point  enough  water  found 
its  way  through  the  fissures  in  the  rock  to  thaw  out  the  sand 
at  the  rock  surface,  and  it  was  necessary  to  again  flood  the 
shaft.  Before  this  was  done,  however,  a  coil  of  pipe  was 


THE  FREEZING   PROCESS  95 

suspended  at  the  rock  surface  and  connected  to  the  freezing 
machine.  This  successfully  stopped  the  leak,  but  six  weeks 
more  were  lost.  The  shaft  was  sealed  to  rock,  and  the  ice 
machine  shut  down  on  June  6,  1889,  after  running  just 
two  hundred  days.  ' 

The  shaft  was  sunk  rectangular  15  ft.  6  in.  X  16  ft.  6  in. 


FIG.  44.  —  Connections  of  Top  of  Freezing  Tube,  Chapin  Shaft 

in  plan,  and  was  lined  with  16  X  16  in.  timber  sets  spaced 
about  4-ft.  centers  on  top,  and  skin  to  skin  at  the  bottom. 
The  frozen  sand  was  blasted  out  with  lime,  black  powder, 
and  finally  dynamite. 

In  1889  the  Mt.  Lookout  Coal  Co.  started  to  sink  two 
shafts  near  Wyoming,  Pa.  The  test  holes  showed  32  ft. 
of  dry  gravel  and  70  ft.  of  quicksand  over  the  rock.  While 
the  first  shaft  was  being  sunk  by  the  pneumatic  process,  an 


96  PRACTICAL  SHAFT  SINKING 

attempt  was  made  to  sink  the  second  by  the  freezing  process. 
Bore  holes  were  put  down  from  5  to  7  ft.  apart  in  a  circle 
around  the  proposed  shaft,  and  cased  through  the  surface 
and  5  ft.  into  the  rock.  A  freezing  mixture  was  then  cir- 
culated in  the  pipe  for  seven  weeks,'  at  which  time  the 
caisson  of  the  first  shaft  reached  rock.  It  was  then  dis- 
covered that  the  rock,  instead  of  being  solid  as  supposed, 
was  fissured  for  18  ft.  below  the  surface.  As  a  large  inflow 
of  water  occurred  in  the  fissures,  the  company  decided  that 
it  would  be  impossible  to  successfully  seal  off  the  water  in  the 
second  shaft  with  the  freezing  tubes  only  5  ft.  in  the  rock, 
and  the  attempt  was  abandoned.  The  shaft  was  then  sunk 
by  the  pneumatic  process,  and  although  some  time  had 
elapsed  between  the  abandonment  of  the  freezing  and  the 
sinking  of  the  caisson,  the  ground  was  found  to  be  still 
frozen  hard. 

The  writer  wishes  to  acknowledge  his  indebtedness  to 
Mr.  J.  Riemer,  from  whose  book,  "  Shaft  Sinking  in  Difficult 
Cases,"*  he  got  many  facts  about  the  sinking  drum  and 
freezing  processes  in  general,  and  a  description  of  the 
European  applications.  The  descriptions  of  the  American 
freezings  were  abstracted  from  the  Transactions  American 
Society  Civil  Engineers  for  June,  1904. 

*  "  Shaft  Sinking  in  Difficult  Cases,"  by  J.  Riemer,  translated  from  the 
German  by  J.  W.  Brough.  J.  B.  Lippincott  &  Co.,  Philadelphia,  1907. 


CHAPTER  VII 

THE  KIND-CHAUDRON  BORING  PROCESS.     CEMENTATION  OF 
WATER-BEARING  FISSURES 

SINKING  in  wet  rock  may  be  accomplished  in  two  ways: 
mechanically,  by  breaking  and  removing  the  rock  under 
water;  by  hand,  by  closing  the  seams  in  the  rock,  thus  pre- 
venting the  inflow  of  water,  or  by  lifting  the  water  as  fast 
as  it  flows  in. 

The  first  plan  can  be  carried  out  in  one  way  only  —  the 
boring  process. 

The  second  is  accomplished  by  the  freezing  process, 
already  described,  and  by  direct  cementation  of  the  fissures 
of  the  rock:  water  is  most  frequently  lifted  by  sinking  pumps 
suspended  in  the  shaft,  although  the  old  Cornish  "spear-rod" 
pump  is  still  sometimes  used  for  large  quantities  of  water. 
A  system  of  water  hoisting  has  also  been  developed. 

The  Boring  Process.  —  The  Kind-Chaudron  boring 
process  has  previously  been  referred  to  as  a  process  devised 
for  sinking  through  rock  measures  containing  such  quanti- 
ties of  water  that  hand  sinking  is  impossible.  It  is  exclu- 
sively a  European  method,  and  so  far  no  shafts  have  been 
bored  in  this  country.  The  process  was  originated  by  M. 
Kind,  a  well  borer,  in  1849.  Between  that  date  and  1854 
he  attempted  to  sink  three  shafts  in  Moselle  and  in  West- 
phalia, but  failed  owing  to  the  inadequacy  of  wooden 
tubbing.  The  scheme  was  then  taken  up  by  M.  J. 
Chaudron,  a  Belgian  engineer  in  France,  and  was  improved 
by  him  to  such  an  extent  that  his  name  is  now  always 
associated  with  that  of  Kind.  Subsequent  improvements 
of  value  have  been  made  by  Riemer  and  others,  and  have 
been  patented  in  Europe;  at  present  the  firm  of  Haniel  & 

97 


98  PRACTICAL  SHAFT  SINKING 

Lueg,  of  Dusseldorf,  controls  many  of  these  patents  and  is 
best  equipped  for  boring  shafts. 

Kind's  original  plan  was  to  bore  the  shaft  in  one  opera- 
tion. The  difficulty  of  collecting  the  broken  rock  made  it 
advisable  to  bore  in  two  stages,  and  a  small  hole,  having  a 
diameter  one-third  to  one-half  that  of  the  finished  shaft, 
is  now  bored  in  advance.  The  muck  in  this  is  removed  by 
a  special  bucket  with  flap  valves  in  the  bottom,  so  arranged 
that  as  the  bucket  is  lowered  the  muck  will  enter.  When 
the  bucket  is  raised  the  valves  close. 

The  small  shaft  serves  as  a  guide  for  the  large  boring 
tool,  as  well  as  for  a  collector  for  muck,  and  it  is  usually 
bored  about  100  ft.  ahead  of  the  large  boring.  The  cutting 
edges  of  the  large  borer  slope  down  toward  the  center;  the 
borings,  therefore,  slide  into  the  small  shaft.  The  large 
tool  thus  has  always  a  clean  surface  to  work  on. 

Boring  is  usually  adopted  as  a  last  resort.  In  regions 
where  large  flows  of  underground  water  are  expected,  the 
shaft  is  sunk  by  hand  until  the  water-bearing  strata  are 
reached  and  is  then  cleared  for  boring.  Care  must  be  taken 
to  keep  the  shaft  free  from  timbers,  permanently  fastened 
pumps,  and  pipes,  etc.,  so  that  when  it  is  flooded  all  impedi- 
ments to  boring  can  be  hoisted  out.  If  this  is  not  done,  it 
will  be  necessary  to  cover  the  bottom  of  the  shaft  with  a 
thick  layer  of  concrete,  deposited  under  water  with  a  grab 
bucket,  in  order  to  permit  the  shaft  to  be  pumped  out  and 
cleared. 

When  boring  is  started,  it  is  carried  through  the  water- 
bearing strata  and  some  distance  into  the  impervious 
ground  beneath.  The  shaft  is  then  lined  with  rings  of  cast- 
iron  tubbing;  the  placing  of  this  tubbing  is  the  most  ingenious 
feature  of  the  boring  process.  (See  Fig.  45.) 

After  the  shaft  has  been  cleaned  out  and  the  boring 
tools  removed,  a  heavy  platform  is  built  over  the  shaft  and 
upon  it  the  "moss  box"  is  erected.  This  consists  of  two 
rings  of  heavy  tubbing,  forming  a  large  stuffing-box  which 
is  filled  with  moss,  and  is  so  designed  that  an  upward  pres- 


THE   KIND-CHAUDRON   BORING   PROCESS 


99 


sure  on  the  lower  ring  will  force  the  moss  out  against  the 
sides  of  the  shaft.  The  moss  is  covered  with  wire  netting 
to  keep  it  in  place,  and  the  diameter  of  the  whole  is  slightly 
less  than  that  of  the  shaft.  On  top  of  the  moss  box  is  bolted 
a  ring,  fitted  with  a  heavy  arched  bulkhead  that  closes  the 


FIG.  45.  —  Tubbing  and  Moss  Box 

shaft.  The  whole  structure  is  then  lifted  by  heavy  jack- 
screws  and,  after  the  platform  is  removed,  is  lowered  into 
the  shaft.  It  is  then  hung  from  beams  placed  across  the 
shaft,  the  jackscrews  are  disconnected  and  withdrawn, 
another  ring  of  tubbing  placed  on  the  beams,  and  the 
jackscrews  reconnected.  The  beams  are  taken  out  and  the 


100  PRACTICAL  SHAFT  SINKING 

ring  is  lowered  into  place  and  bolted  up.  The  process  is 
then  repeated. 

Since  each  ring  of  tubbing  weighs  less  than  the  water  it 
displaces,  after  enough  have  been  added  the  whole  column 
of  tubbing  will  float.  The  hanging  rods  are  then  dispensed 
with  and,  as  each  ring  is  bolted  on,  the  tubbing  is  sunk  by 
admitting  water.  This  is  continued  until  the  moss  box 
reaches  the  bottom  of  the  shaft.  The  whole  column  is 
then  allowed  to  fill  with  water,  and,  all  buoyancy  being  re- 
moved, the  entire  weight  of  the  tubbing  serves  to  crush  the 
moss  outward  against  the  sides  of  the  shaft.  A  water-tight 
joint  is  thus  made  at  the  bottom. 

The  space  between  the  tubbing  and  the  sides  of  the  shaft 
is  then  filled  with  concrete,  and,  after  this  has  set  long 
enough  to  thoroughly  harden,  the  shaft  is  pumped  out  and 
the  bulkhead  removed.  The  concrete  is  usually  placed 
with  small  flat  buckets  lowered  with  ropes. 

The  boring  tools,  borers  or  " trepans,"  Fig.  46,  are  made 
of  cast  steel  provided  with  inserted  teeth,  and  weigh  about 
10  tons  for  the  small  and  20  tons  for  the  large  tool.  They 
are  suspended  by  heavy  timber  rods  from  a  walking  beam 
operated  by  a  single  large,  vertical  steam  cylinder.  Between 
the  walking  beam  and  the  rods  a  chain  connection  is  pro- 
vided, by  means  of  which  the  borers  can  be  lowered  as  the 
hole  deepens.  At  the  "bore  master's"  platform  on  top 
of  the  shaft  there  is  a  swivel  connection  and  a  long  lever  for 
rotating  the  borer  slightly  between  blows. 

The  head-house  is  a  tall  structure  with  two  wings  in 
which  are  stored  the  various  tools.  All  are  hung  from  small 
trucks  which  can  be  readily  run  out  over  the  shaft.  Two 
hoist  engines  and  cables  are  provided,  one  for  the  bucket 
and  the  other  for  the  borers  and  tools. 

Considerable  difficulty  is  encountered  in  boring  through 
fissured  ground  and  through  strata  of  soft  material.  In  the 
first  case  the  teeth  of  the  borers  are  liable  to  breakage;  in 
the  second,  large  masses  of  material  fall  out  of  the  sides  of 
the  shaft,  sometimes  burying  the  borer.  A  special  tool  has 


CEMENTATION  OF  WATER-BEARING  FISSURES 


101 


been  devised  for  fishing  out  broken  teeth,  large  pieces  of 
rock,  etc.  A  falling  in  of  material  is  prevented  by  sheet- 
iron  cylinders,  lowered  from  the  top  and  suspended  by  flat 
ropes.  These  cylinders  have  no  hydrostatic  pressure  to 
resist,  hence  need  not  be  heavy.  They  are  built  of  f-in. 
plate,  in  lengths  up  to  60  ft. 


FIG.  46.  —  Cross-section  of  Boring  Tower 

It  has  not  been  found  practicable  to  use  segmental 
tubbing  rings  for  lining  bored  shafts,  owing  to  the  difficulty 
of  making  the  vertical  joints  water-tight.  The  maximum 
diameter  at  present,  therefore,  is  limited  by  the  size  of  the 
single  ring  which  can  be  transported  —  about  14  ft.  diameter. 
The  boring  is  made  18  in.  to  3  ft.  larger,  depending  upon  the 
character  of  the  ground.  Ordinarily  the  small  borer  is 
about  8  ft.  across,  and  the  large  one  15^  ft.  for  14-ft.  tub- 
bing, but  if  much  bad  ground  is  expected  and  the  use  of 


102  PRACTICAL  SHAFT  SINKING 

several  sheet-iron  cylinders  is  contemplated,  the  large  borer 
is  made  16|  ft.  or  17  ft.  across  to  start  with. 

The  speed  made  in  boring  shafts  has  varied  so  greatly 
that  it  is  hard  to  give  definite  figures.  Under  favorable 
conditions  the  small  tool  will  advance  20  in.  per  day,  and 
the  large  one  7  or  8  in.  If  to  the  time  required  for  actual 
boring  is  added  that  taken  for  lowering  sheet-iron  cylinders 
and  tubbing,  it  will  be  seen  that  an  average  progress  of  8  to 
10  ft.  per  month  is  all  that  can  be  expected.  The  costs  of 
the  boring  process  are,  in  consequence,  exceedingly  high, 
but  for  the  very  difficult  sinking  conditions  obtaining  in 
some  parts  of  Europe  it  is  the  only  process  that  is  unfailingly 
successful.* 

Direct  Cementation.  —  The  injection  of  cement  grout 
under  pressure  into  fissured  rock  has  been  attempted  only 
recently.  Immediately  around  the  proposed  shaft  a  num- 
ber of  holes  are  drilled  through  the  fissured  rock  into  the 
solid  measures  beneath,  and  grout  is  forced  into  them  until 
all  crevices  are  filled.  It  is  then  allowed  to  set,  and,  if  the 
work  has  been  properly  done,  sinking  can  be  continued  in 
the  dry. 

The  water-bearing  fissures  can  best  be  located  by  using 
core  drills  rather  than  percussion  drills  in  boring  the  holes. 
If  only  one  crevice  exists,  the  grout  will  flow  directly- into 
it;  if,  however,  the  rock  is  fissured  for  some  distance,  while 
grout  is  flowing  into  the  upper  cracks,  the  hole  beneath  them 
may  become  blocked  before  the  lower  cracks  are  filled.  In 
this  case  the  water  will  not  be  completely  shut  off. 

Up  to  the  present  time  the  blocking  of  the  fissures  in  any 
considerable  depth  of  rock  has  only  been  accomplished  by 
successive  cementations.  This  was  done  in  the  two  cases 
described  below.  The  writer  believes,  however,  that  if 
the  holes  are  bored  entirely  through  the  wet  rock,  a  method 
can  be  devised  for  filling  the  cracks  from  the  bottom  up. 
A  plan  that  might  be  worked  out  would  be  to  case  the  holes 

*For  detailed  infomation  on  this  process  see:  "  Shaft  Sinking  in  Difficult 
Cases,"  by  J.  Riemer. 


CEMENTATION  OF  WATER-BEARING  FISSURES          103 

with  flush-joint  pipes  to  the  bottom,  then  to  gradually 
withdraw  the  pipes  as  the  grout  is  pumped  in.  The  grout- 
ing apparatus  would  have  to  be  so  arranged  as  to  permit 
quick  disconnection  and  reconnection  upon  the  removal 
of  each  length  of  pipe,  to  avoid  the  possibility  of  the  grout 
setting  in  the  pipe  while  the  flow  is  interrupted. 

The  following  account  of  the  cementation  of  a  shaft 
sunk  by  the  Mining  Society  of  Lens  is  an  abstract  from  an 
account  published  by  C.  Dinoire  in  Volume  XXXI  of  the 
Transactions  of  the  Institute  of  Mining  Engineers  (English) : 

The  Mining  Society  of  Lens  decided  in  October,  1904, 
to  sink  two  shafts,  one  by  the  freezing  method  and  one  by 
hand.  The  water  encountered  in  the  second  exceeded  the 
expectations  to  such  an  extent  that  it  could  not  be  pumped 
and  the  feeders  had  to  be  stopped  by  direct  cementation. 

This  shaft  was  sunk  to  a  depth  of  166  ft.  before  an  inflow 
of  2200  gallons  per  minute  made  cementation  necessary. 
The  pumps  were  worked  very  hard  to  hold  the  water  down 
as  low  as  possible,  and  two  lines  of  2-in.  pipe,  extending 
from  the  surface  to  the  bottom  of  the  shaft,  were  installed. 
As  some  of  the  water  came  in  through  a  nearly  vertical 
fissure  in  the  shaft  bottom,  and  some  through  a  horizontal 
seam,  one  of  the  pipes  was  driven  into  the  fissures  for  9  ft. 
and  the  other  was  terminated  opposite  to  the  seam,  Fig.  46. 
Four  8-in.  bore  holes,  197  ft.  deep,  were  then  put  down  out- 
side the  shaft  area.  They  were  spaced  evenly  13  ft.  from 
the  circumference  of  the  shaft.  At  a  depth  of  190  ft.  they 
passed  through  a  bed  of  very  seamy  rock.  A  hand  pump 
was  put  on  each  of  these  holes  in  order  to  pump  out  all 
sand  and  mud  caused  by  boring. 

Three  hundred  and  sixty  sacks  of  cement  mixed  as  a 
thin  grout  were  then  run  into  the  pipe  which  terminated 
opposite  the  horizontal  seam.  This  took  seven  hours;  the 
grout  mixer  was  then  connected  to  the  other  pipe  to  give 
the  grout  in  the  bottom  of  the  shaft  a  chance  to  set.  Two 
hundred  and  ninety  sacks  were  run  into  this  pipe  in  three 
hours,  when  the  lower  end  became  blocked.  Ninety  sacks 


104 


PRACTICAL  SHAFT   SINKING 


were  then  run  into  the  first  pipe,  completely  filling  it,  after 
which  the  bore  holes  were  filled.  This  required  475  sacks. 
The  pumps  on  the  bore  holes  were  worked  continuously 
while  the  pipes  were  being  grouted. 

The  grout  was  allowed  to  set  for  nineteen  days  before 
the  water  was  pumped  out  of  the  shaft.  It  was  then  found 
that  the  leak  was  completely  closed  and  a  2|-ft.  layer  of 


FIG.  47.  —  Location  of  Pipes  and  Fissures  in  Lens  Shaft 

grout  had  formed  on  the  bottom.     The  rest  of  the  grout, 
amounting  to   282  cu.   ft.,  had  run  into  the  fissures. 

The  bed  of  grout  was  sunk  through  very  carefully,  and 
the  shaft  deepened  to  170  ft.  and  lined  to  the  bottom.  At 
185  ft.  a  second  inflow  was  encountered,  the  water  breaking 
in  through  the  vertical  fissure  referred  to  above.  Two  more 
grout  pipes  were  inserted,  their  lower  ends  being  driven 
3  and  6  ft.  into  the  fissure.  Nine  hundred  sacks  of  cement 


CEMENTATION  OF  WATER-BEARING  FISSURES          105 

were  run  into  the  longer  pipe  in  nine  hours,  when  the  lower 
ends  of  both  pipes  became  closed. 

After  twenty-eight  days  the  water  was  pumped  out  and 
sinking  was  resumed.  The  large  fissure  was  found  to  be 
this  time  thoroughly  cemented,  and  further  sinking  and  lining 
was  carried  on  without  particular  difficulty. 

It  was  found  in  the  first  cementation,  where  the  mixer 
discharged  directly  into  2-in.  pipes,  that  considerable  air 
was  carried  down  with  the  grout.  This  hindered  the  flow 
very  greatly,  so  that  in  the  second  cementation  the  grout 
pipes  were  made  lj  in.  diameter  above  water  and  2f  in. 
under  water,  a  high  narrow  tank  was  connected  with  the 
top  of  the  grout  pipe,  and  a  valve  was  provided  to  regulate 
the  flow  from  the  tank,  which  was  kept  full.  The  mixer 
discharged  into  the  tank  through  a  trough,  with  gratings 
for  removing  air. 

The  more  important  conclusions  reached  were: 

1.  Fissures  in  rock  can  be  cemented  through  pipes  or 
bore  holes. 

2.  Sand,    marl,    boulders,    clay,   and   slime   cannot   be 
cemented. 

3.  Thin  beds  presenting  continuous  openings  can  be 
cemented  by  pumping  from  bore  holes  outside  the  shaft. 

Direct  cementation  has  also  been  applied  at  the  Anzin 
and  the  Bethune  collieries  in  Europe. 

Direct  cementation  has  only  been  attempted  in  America 
in  one  instance.  No.  4  shaft  on  the  Rondout  siphon  of 
the  Catskill  Aqueduct  encountered  a  flow  of  water  of  750 
gallons  per  minute  at  a  depth  of  270  ft.  The  shaft  is  rect- 
angular with  three  compartments,  measures  only  8  ft.  4  in. 
X  20  ft.  4  in.  in  the  clear,  and  thus  affords  very  little  oppor- 
tunity for  handling  large  pumps.  In  addition  to  this  fact 
the  water  was  strongly  charged  with  sulphureted  hydrogen 
gas,  which  acted  very  painfully  upon  the  eyes  of  the  sinkers. 

After  no  progress  had  been  made  for  several  weeks,  it 
was  decided  to  try  grouting.  Most  of  the  water  came 
directly  out  of  the  bottom.  The  water  level  was  held  within 


106  PRACTICAL  SHAFT  SINKING 

about  5  ft.  of  the  bottom  and  a  number  of  holes  from  10  to 
18  ft.  deep  were  drilled  with  rock  drills.  Two-inch  pipes 
were  connected  to  these  holes,  and  about  1500  sacks  of 
cement,  mixed  as  neat  grout,  were  forced  into  them.  This 
reduced  the  flow  of  water  in  the  bottom  from  525  to  about 
50  gallons  per  minute. 

In  order  to  block  the  fissures  below  the  shaft  bottom,  a 
diamond  drill  was  set  up  on  a  platform  at  the  top  of  the  shaft, 
and  six  holes  were  drilled,  each  about  95  ft.  deep.  A  column 
of  3-in.  pipe  extending  from  the  top  to  the  bottom  of  the 
shaft  served  in  each  case  as  a  guide  for  the  drill  rod.  These 
holes  passed  entirely  through  the  water-bearing  rock,  which 
consisted  of  a  badly  fissured  sandstone,  and  penetrated  an 
impervious  stratum  of  grit  beneath.  Two-inch  grout  pipes 
were  now  connected  to  these  diamond-drill  holes  and  183 
sacks  of  cement  were  forced  into  them.  This  entirely  cut 
off  the  water  in  the  bottom. 

It  was  found  upon  pumping  out  the  water  that  all  the 
upper  fissures  were  completely  filled.  When  sinking  was 
resumed,  however,  it  was  further  found  that  the  drill  holes 
had  become  blocked  before  the  lower  fissures  were  completely 
filled,  and  the  water,  therefore,  was  not  altogether  shut 
off.  After  the  cementation  it  was,  nevertheless,  possible 
to  handle  the  water  with  pumps  and  sink  about  10  ft.  a 
week  in  the  ordinary  way. 

In  filling  all  the  holes  the  grout  pipes  were  carried  to  the 
top  of  the  shaft.  The  grout  was  mixed  and  fed  into  a  tank 
attached  to  the  top  of  each  pipe.  When  no  more  grout 
would  flow  by  gravity,  the  opening  in  the  tank  was  closed, 
and  air  pressure  applied  on  top  of  the  liquid  grout.  The 
pressure  ranged  from  80  Ibs.  to  a  maximum  of  300  Ibs.  per 
square  inch,  which  last  was  obtained  by  means  of  a  special 
air  compressor. 

NOTE. — For  further  information  about  methods  of  grouting  see  Ap- 
pendix A. 


CHAPTER  VIII 

LIFTING    WATER.         HORIZONTAL   vs.    VERTICAL   PUMPS. 
HANDLING  PUMPS  IN  SHAFT.     CORNISH  PUMPS 

MODERATE  quantities  of  water  are  ordinarily  raised 
from  the  bottom  of  a  sinking  shaft  by  one  or  more  pumps, 
suspended  or  supported  in  the  shaft  just  above  the  bottom, 
and  lowered  from  time  to  time  as  the  shaft  is  deepened. 
The  usual  motive  power  is  steam  or  compressed  air;  electric 
pumps  are  being  developed,  but  so  far  have  not  been  suc- 
cessful. 

The  work  that  a  sinking  pump  is  called  upon  to  do  is 
exceedingly  arduous.  It  must  first  of  all  be  reliable;  it 
must  run  on  gritty  water  or  a  mixture  of  water  and  air,  or 
sometimes  on  air  alone  for  a  while  without  injury.  The 
valves,  packing,  and  wearing  parts  must  be  readily  acces- 
sible. It  must  occupy  a  minimum  space  in  the  shaft,  and 
at  the  same  time  be  strong  and  heavy  enough  to  endure 
collisions  with  the  sides  in  hoisting  and  lowering,  and  blows 
from  flying  fragments  of  rock. 

Any  one  in  this  country  who  has  tried  to  sink  a  wet 
shaft  is  more  or  less  familiar  with  the  features  of  the  leading 
American  sinking  pumps.  Of  these  the  Cameron  is  the  best 
known,  and  the  Cameron  pattern,  Fig.  48,  now  manufac- 
tured by  a  number  of  firms,  certainly  comes  nearest  to  filling 
the  requirements.  This  is  built  in  both  vertical  and  hori- 
zontal piston  and  outside-packed  plunger  styles,  and  is 
marked  by  the  absence  of  outside  valve  gear,  by  the  thick- 
ness and  strength  of  the  castings,  and  by  the  accessibility 
of  the  water  valves  and  packing. 

The  manufacturers  as  a  rule  recommend  the  vertical 
pump  for  sinking,  but  the  writer's  experience  has  taught 

107 


108 


PRACTICAL  SHAFT  SINKING 


him  otherwise.  When  there  is  only  a  slight  inflow  of  water, 
a  small  vertical  pump  with  hose  connections  is  certainly 
very  easy  to  hang  in  the  shaft;  on  the  other  hand,  a  hori- 
zontal pump,  with  a  capacity  of  150  gallons  per  minute 


FIG.  48.  —  Vertical  Sinking  Pump 

or  less,  will  work  perfectly  when  hung  with  a  bridle.  Where 
there  is  a  considerable  flow  of  water,  and  where  the  shaft  is 
large  enough  to  accommodate  a  horizontal  pump  or  pumps 
of  the  required  capacity,  this  is  the  type  to  use.  Where  the 


LIFTING  WATER  109 

shaft  is  so  small  that  there  is  not  room  in  it  for  enough 
horizontal  pumps  to  take  care  of  the  water,  a  vertical 
pump  is  of  course  a  necessity. 

The  writer's  reasons  for  preferring  the  horizontal  pump 
are: 

1.  The  horizontal  pump  is  lighter  than  a  vertical  of  the 
same  capacity. 

2.  In  the  larger  sizes  the  horizontal  pump  is  much  more 
accessible,  since  a  man  can  walk  around  it  and  work  at  any 
part  of  it  on  a  level  platform.     It  is  necessary  to  climb  6  or 
8  ft.  to  get  from  the  water  end  to  the  steam  end  of  a  big 
vertical  pump. 

3.  By  providing  a  proper  bridle,  the  horizontal  pump 
can  be  hooked  on  to  and  lifted  as  easily  as  the  vertical. 

4.  For  a  pump  discharging  over  300  gallons  per  minute, 
the  recoil  at  every  stroke  is  so  severe  that  hose  or  other 
flexible  connections  will  not  stand  when  the  pump  is  hung 
freely.     This  applies  to  the  American  style  of  vertical  sinking 
pump,  where  the  center  line  of  neither  suction  nor  discharge 
coincides  with  the  center  of  suspension,  as  well  as  to  a  hori- 
zontal pump.     In  both  cases  it  is  necessary  to  furnish  rigid 
support,  and  it  is  as  easy  to  set  two  hitch  timbers  in  a  hori- 
zontal plane  for  a  horizontal  pump  as  it  is  to  set  them  in  a 
vertical  plane  for  a  vertical  pump. 

In  sinking  several  shafts  in  which  the  flow  of  water  in 
the  bottom  varied  from  800  to  1500  gallons  per  minute,  the 
writer  has  obtained  the  best  results  by  working  along  the 
following  lines: 

Provide  an  absolutely  reliable  boiler  plant  of  ample 
capacity.  Sinking  pumps  are  frightfully  uneconomical, 
requiring  from  200  to  250  Ibs.  of  steam  per  actual  horse- 
power hour;  a  shaft  in  which  1000  gallons  per  minute  must 

,     ,.f.    ,  onn  ,.       .„   AU      .  .      1000  X  84  X  300  ft. 

be  lifted  300  ft.  will,  therefore,  require — — 

33,000 
200 
X =  505  horse-power  of  boilers  for  pumping  alone. 

OvJ 

Put  in  a  water  ring,  Fig.  49,  and  a  stationary  pump 


110 


PRACTICAL  SHAFT  SINKING 


wherever  it  is  possible  to  reduce  the  water  in  the  bottom 
by  so  doing.  When  the  water-bearing  stratum  has  been 
sunk  through,  open  up  a  "lodgment"  or  reservoir  in  one 
end  of  the  shaft  and  install  a  compound  pump. 

For  handling  water  in  the  bottom  provide  at  least  50 
per  cent,   extra  pumping    capacity;   say,   four   350-gallon 


FIG.  49.  —  Section  of  Water  Ring  for  Timbered  Shaft 

pumps  for  900  gallons  per  minute,  two  in  each  end  of  the 
shaft.  Set  bearing  timbers,  Fig.  50,  as  close  to  the  bottom 
as  is  safe,  and  lower  pumps  alternately  10  ft.  at  a  time, 
keeping  10-ft.  and  20-ft.  flanged  lengths  of  steam,  exhaust, 
and  water  pipes  ready  to  put  on.  If  each  new  set  of  timbers 
is  placed  10  ft.  from  the  bottom  of  the  shaft,  the  maximum 
suction  lift  at  any  time  will  thus  be  20  ft.  Make  swinging 
joints  on  all  pipe  lines  at  the  pumps,  to  take  care  of  vibra- 
tion and  expansion  and  of  variations  in  the  spacing  of  the 


LIFTING   WATER 


111 


hitch  timbers.     Do  not  attempt  to  lift  pumps  when  blasting; 
remove  suction  hose  only  and  shoot  carefully. 

Provide  an  independent  column  pipe  for  each  pump,  and 
keep  the  pipes  in  the  hoistway.  Main  steam  and  exhaust 
lines  should  be  kept  in  one  end  of  the  shaft  for  the  greater 
part  of  their  length  in  order  to  create  an  upward  current  of 


FIG.  50.  —  Arrangement  of  Pumps  for  Sinking 

air  in  that  end  and  assist  ventilation.  Handle  pumps  with 
independent  engine,  arranged  to  hoist  from  any  compart- 
ment. Clamp  pipes  to  timber  every  three  lengths. 

Too  much  care  cannot  be  taken  to  obtain  tight  joints  in 
all  pipes,  to  fasten  them  securely  to  the  timbers,  to  keep  the 
pumps  in  good  repair  and  plenty  of  spare  parts  (valves, 


112  PRACTICAL  SHAFT  SINKING 

stems,  packing,  etc.)  on  hand,  and  to  keep  the  follower 
bolts  and  nuts  tight.  It  is  advisable  to  fasten  these  securely 
with  cotter  pins,  or  by  drilling  through  the  heads  and  wiring 
them  together. 

Pumps  will  run  satisfactorily  on  compressed  air  if  the 
air  is  reheated  sufficiently  to  prevent  freezing.  As  a  rule, 
the  cost  of  the  air  plant  makes  it  imperative  to  use  steam 
on  the  pumps.  Both  steam  pipes  and  exhaust  pipes  should 
be  lagged  with  some  water-proof  covering. 

At  best,  many  delays  will  occur  when  water  is  handled. 
For  a  shaft  making  800  to  900  gallons  per  minute,  15  to  20  ft. 
per  month  is  good  progress;  the  labor  cost  may  easily  reach 
$150  to  -1200  per  foot.  Sinking  with  insufficient  machinery 
is  impossible. 

In  this  country  water  is  usually  regarded  as  an  unfor- 
tunate accident,  and  when  encountered  is  met  by  begrudged 
additions  to  the  plant.  As  a  result  much  time  and  money 
are  wasted.  In  Europe,  on  the  other  hand,  the  mining  dis- 
tricts have  been  more  fully  explored  and  developed,  and  the 
position  of  water-bearing  strata  is  known.  Preparations 
are  made  in  advance  for  handling  large  quantities  of  water. 
The  shafts  are  circular  and  are  lined  in  sections  as  the  sink- 
ing proceeds  with  brick  or  water-tight  iron  tubbing;  their 
large  diameter  and  freedom  from  cross-braces  make  it  pos- 
sible to  use  more  powerful  pumps.  An  actual  measured 
flow  of  from  2000  to  2500  gallons  per  minute  in  the  shaft 
bottom  is  as  much  as  has  been  successfully  taken  care  of  in 
America.  In  the  two  English  cases  cited  below  two  to  three 
times  this  much  water  was  pumped. 

In  the  Transactions  of  the  Federated  Institute  of  Mining 
Engineers,  Volume  III,  page  513,  Mr.  W.  H.  Chambers 
describes  the  sinking  of  two  shafts  at  Conisboro,  Yorkshire. 
Eight  30  ft.  X  7  ft.  6  in.  Lancashire  boilers  were  installed,  and 
sinking  was  then  commenced  in  both  shafts  simultaneously. 
The  water  was  handled  by  pulsometers  to  a  depth  of  156  ft., 
when  an  inflow  of  water  occurred  that  made  it  necessary 
to  put  in  very  much  more  powerful  pumps. 


LIFTING   WATER  113 

The  sinking  pumps,  Fig.  51,  were  made  by  Baily  &  Co., 
of  Salford,  and  were  designed  to  run  suspended  in  the  shaft 
without  other  support  than  two  suspension  ropes  which  also 
carried  all  pipe  lines.  A  telescopic  suction  pipe  is  provided 
instead  of  suction  hose,  and  the  axes  of  both  suction  and  dis- 
charge pipes  coincide  with  the  axis  of  suspension.  All 
vibrations  caused  by  the  strokes  of  the  pump  are,  therefore, 
vertical  and  cause  no  dangerous  sideways  motion  in  the  pipe 
lines.  The  pump  itself  " consists  of  three  hollow  plungers; 
the  upper  pair  are  stationary  and  over  them  slide  barrels 
which  are  connected  to  the  steam  piston.  The  third  barrel 
is  secured,  together  with  the  pair  of  stationary  plungers,  to 
the  steam  cylinder  by  means  of  connecting  rods."  The 
pump  is  thus  what  is  here  known  as  the  differential  plun- 
ger type.  Two  discharge  pipes  are  led  from  the  top  of  the 
upper  stationary  plungers  alongside  the  steam  cylinder, 
and  are  joined  above  it  by  a  tee  from  which  the  column 
pipe  rises. 

The  two  suspension  ropes  are  led  over  pulleys  at  the 
shafthead  to  the  drum  of  a  hoisting  (or  capstan)  engine,  and 
the  pump  and  all  piping  are  hoisted  and  lowered  together. 
A  telescopic  joint  is  put  on  the  steam  pipe  at  the  shaft  head, 
the  discharge  pipe  is  turned  sideways  over  a  trough,  and  the 
exhaust  pipe  stands  straight  up.  The  arrangement  of  pump 
and  pipes  in  the  shaft  is  shown  in  Fig.  52  and  the  method 
of  supporting  the  pipes  is  shown  in  horizontal  section  in 
Fig.  52a. 

Six  pumps  of  this  type,  each  with  a  capacity  of  50,000  to 
70,000  Imperial  gallons  per  hour  at  35  strokes  per  minute, 
were  needed  to  sink  the  shafts  to  a  depth  of  300  ft.  In  this 
depth  the  maximum  quantity  of  water  discharged  from 
both  shafts  amounted  to  6600  gallons  per  minute,  although 
the  tubbing  was  carried  along  with  the  sinking.  Two  more- 
pumps  were  then  obtained,  and  the  eight  were  arranged  to 
lift  the  water  in  two  stages,  as  300  ft.  was  the  maximum  lift 
of  each  pump.  The  water  was  finally  shut  off  at  a  depth  of 
395  ft.  in  one  shaft  and  369  ft.  in  the  other. 


LIFTING  WATER 


115 


Mr.  Chambers  describes  the  operation  of  the  pumps  as 
follows : 

"As  the  sinking  progressed,  after  the  suction  pipe  was 


FIG.  52 


FIG.  52a 


drawn  out  to  its  full  extent,  the  pump  with  the  columns  of 
pipe  was  lowered  by  running  the  ropes  off  the  capstan,  and 
exhaust  and  water  pipes  were  built  as  required  on  top;  the 


116  PRACTICAL  SHAFT  SINKING 

steam  pipe,  after  being  drawn  its  full  length  out  of  the 
stuffing-box,  was  pushed  back  and  another  length  inserted. 

"A  stop  valve  and  a  lubricator  were  placed  in  the  fixed 
steam  pipe  on  the  surface.  A  lad  was  in  charge  to  regulate 
the  supply  of  steam  .as  required,  he  being  in  communication 
with  the  sinkers  in  the  shaft  by  means  of  a  signal  bell.  The 
speed  of  the  pumps  was  thus  controlled  and  lubrication 
effected  without  any  one  being  in  the  shaft  for  these  pur- 
poses." 

Perhaps  the  most  remarkable  achievement  in  the  line  of 
wet-shaft  sinking  that  was  ever  accomplished  was  the  sink- 
ing at  the  Horden  colliery  in  Southeast  Durhamshire. 
This  work  is  described  at  length  in  a  very  interesting  paper 
by  Mr.  J.  J.  Prest,  the  engineer  in  charge.  The  paper  is 
published  in  the  Proceedings  of  the  Institution  of  Civil 
Engineers,  Volume  CLXXIII,  Part  3. 

At  this  colliery  three  shafts  were  sunk,  two  of  them 
simultaneously;  the  third  was  put  down  to  the  level  at  which 
the  greatest  flow  of  water  occurred,  and  there  stopped  until 
the  other  two  reached  the  coal  measures.  Mr.  Prest,  after 
careful  consideration,  determined  to  use  the  old-style  Cornish 
pump  and  installed  a  very  remarkable  plant,  comprising 
over  3000  boiler  horse-power  and  no  less  than  four  sets  of 
30-in.  bore  by  6-ft.  stroke  pumps.  Each  set  consisted  of  a 
pair  of  pump  cylinders  hung  so  as  to  balance  each  other,  and 
capable  of  being  arranged  as  a  high-and  a  low-lift  set.  The 
pumps  were  driven  by  the  permanent  hoisting  engines, 
provided  with  an  extra  jack-shaft  and  gearing. 

The  maximum  quantity  of  water  handled  simultaneously 
from  all  three  shafts  amounted  to  9230  Imperial  gallons  per 
minute,  and  the  maximum  from  one  shaft  to  6310  gallons 
per  minute,  this  quantity  being  pumped  from  a  depth  of 
300  ft.  The  shafts  reached  an  average  depth  of  about  540 
ft.  before  the  coal  measures  were  reached  and  it  was  finally 
possible  to  tub  back  the  water. 

A  system  has  been  developed  in  England  for  hoisting 
water  from  a  sinking  shaft  without  the  use  of  any  high- 


LIFTING  WATER  117 

pressure  pumps.  It  is  known  as  the  Tomson  water-winding 
process.  Mr.  Tomson  puts  in  his  permanent  hoisting 
engine,  places  guides  in  the  shaft  as  the  sinking  proceeds, 
and  uses  large  tanks  for  lifting  the  water.  The  tanks  are 
filled  by  low-pressure  pumps  driven  by  compressed  air  and 
attached  to  the  tanks.  This  system  has  been  very  success- 
ful in  some  instances,  but  has  not  been  used  where  the  quan- 
tities of  water  were  as  great  as  those  at  Conisboro  or  Horden. 


CHAPTER  IX 

SHAFT  LININGS 

SHAFTS  are  usually  lined;  either  in  order  to  exclude 
water,  or  to  support  the  sides  and  prevent  the  falling  of 
fragments  of  rock. 

The  most  common  lining  material  —  in  fact  until 
recently  almost  the  only  lining  material  —  used  for  American 
mine  shafts  is  timber,  ordinarily  framed  in  square  sets  and 
lagged  with  plank.  Such  a  lining  cannot  be  made  water- 
tight and  acts  only  as  a  support  or  shield.  Wooden  caissons 
and  coffers  used  in  bad  surface  ground  are  of  course  built 
to  exclude  water,  but  this  construction  is  not  feasible  at 
considerable  depths. 

European  shafts  are  usually  lined  with  brick  walls, 
built  upon  cast-iron  curb  rings  set  into  the  sides  of  the  shaft 
at  intervals.  The  circular  and  elliptical  sections  universal 
in  Europe  are  in  fact  accounted  for  by  the  necessity  for  arch 
action  in  a  brick  lining.  The  walls  are  not  designed  to 
absolutely  exclude  water,  but  to  lead  it  to  water  rings  at  the 
curbs  and  prevent  it  from  dripping  in  the  shaft.  A  brick 
lining  is  fire-proof  and  durable. 

Where  it  is  desired  to  block  back  large  feeders  of  water, 
cast-iron  tubbing  is  used.  This  has  already  been  referred 
to  in  connection  with  the  Boring  Process.  Two  styles  of  it 
are  used  in  Europe,  known  respectively  as  English  and  Ger- 
man tubbing;  the  writer  knows  of  no  case  where  it  has  been 
used  for  a  mining  shaft  in  this  country,  but  miles  of  suba- 
queous tunnel  around  New  York  City  are  lined  with  bolted 
cast-iron  segments. 

In  the  last  decade,  concrete  has  come  to  the  front  as  a 
material  for  lining  shafts.  It  is  gradually  being  realized 

118 


SHAFT  LININGS  119 

that,  when  properly  handled,  concrete  can  be  made  as  water 
tight  as  iron;  it  is  stronger  than  brick  and  as  durable,  and  is 
cheaper  than  either  iron  or  brick.  The  first  shaft  in  America 
to  be  entirely  lined  with  concrete  was  sunk  by  the  U.  S.  Coal 
and  Coke  Co.  at  Tug  River,  West  Virginia,  in  1903.  Since 
then  this  construction  has  been  adopted  for  a  dozen  shafts 
or  more/ 

Of  the  various  kinds  of  shaft  lining,  timber  is  the  easiest 
to  place  and  in  America  is  still  the  cheapest  in  first  cost. 
This  advantage,  however,  diminishes  every  year;  good  tim- 
ber is  scarce  and  dear  to-day,  and  ten  or  fifteen  years  hence 
—  the  life  of  a  timber  lining  —  will  be  scarcer  and  dearer.  A 
timbered  shaft  in  a  mine  whose  life  is  expected  to  exceed  the 
life  of  the  original  timber  is  thus  a  very  doubtful  investment. 

Considerations  of  safety  present  a  stronger  argument 
against  the  use  of  timber  in  coal  mine  shafts.  A  severe 
explosion  in  the  mine  will  wreck  the  lining  and  fill  .the  shaft 
with  twisted  timber,  thus  cutting  off  all  hope  of  escape  or 
rescue  from  the  men  imprisoned  below.  This  danger  has 
long  been  avoided  in  Europe  by  the  use  of  walled  shafts, 
and  before  many  years  public  sentiment  in  America  will 
demand  that  it  be  avoided  here.  A  Pennsylvania  law  now 
on  the  statute  books  prohibits  the  use  of  wood  in  permanent 
tipples  or  breakers  within  200  ft.  of  the  shaft  head;  why 
should  not  this  law  be  logically  extended  to  cover  timber 
in  the  shafts  itself? 

In  ore  shafts  and  construction  shafts  the  same  objection 
to  timber  does  not  apply,  although  the  possibility  of  fire 
must  be  considered.  This  is  not  often  a  serious  danger  as 
the  timbers  are  usually  so  wet  that  nothing  short  of  an 
explosion  could  ignite  them.  A  system  of  reinforced  con- 
crete " timbers"  which  has  been  proposed  and  patented 
is  intended  to  retain  the  advantages  of  a  rectangular  tim- 
bered shaft,  and  at  the  same  time  provide  a  fire-proof 
lining  that  is  easily  placed.  It  has  never  been  used. 

Timbering.  Buntons.  —  In  rock  that  is  hard  enough  to 
stand  without  support  and  is  not  affected  by  frost  and 


120 


PRACTICAL  SHAFT  SINKING 


moisture,  the  cheapest  construction  is  an  unlined  rectan- 
gular shaft  with  several  vertical  rows  of  buntons  to  which 
the  guides  and  ladders  are  fastened.  These  buntons  cor- 
respond to  the  buntons  and  end  plates  of  a  square-framed 
set  of  timbers,  and  are  set  into  pockets  cut  in  the  rock. 


- 
Compartment 


.8"*/o'/?i.ntj  Umbers 


*5  fbr  Guides  ^ 


DetacL    of  corner 


Other  Method:,  of  securing 

FIG.  53.  — Timber  Sketches 

They  must  be  set  correctly  in  parallel  vertical  planes, 
and  should  be  level  and  in  line  horizontally,  but  absolute 
accuracy  in  this  respect  is  not  essential.  The  pockets  are 
known  as  "hitches,"  the  "box  hitch"  is  cut  square;  the 
"drop  hitch"  is  cut  with  the  top  sloping  back  so  that  after 
one  end  of  the  bunton  is  placed  in  the  box  hitch  the  other 


SHAFT  LININGS 


121 


may  be  dropped  into  place.  The  buntons  are  usually  set 
on  5-ft.  centers  vertically,  so  that  workmen  standing  on 
a  platform  on  one  row  of  timbers  can  cut  the  hitches  for 
the  next  row  without  scaffolding.  (See  Fig.  54.) 

After  the  buntons  are  placed  in  the  hitches  they  are 
secured  and  held  in  line  by  oak  wedges  driven  in  tight. 
The  hitches  are  cut  just  enough  larger  than  the  timber  to 
allow  for  wedging.  The  depth  depends  upon  the  rock; 
specifications  usually  call  for  12  to  18-in.  hitches,  but,  as  a 
matter  of  fact,  in  any  rock  hard  enough  to  stand  without 
support  a  4-in.  hitch  will  break  the  timber. 


FIG.  54.  —  Shaft  Timbered  with  Buntons  Only 

Hitches  are  usually  cut  by  hand  with  hammer  and  bull 
point,  but  pneumatic  hammers  can  be  used  to  advantage 
in  hard  rock.  When  working  by  hand  a  pair  of  men  (ham- 
mersman  and  holder)  should  finish  a  pair  of  hitches  in  an 
eight-hour  shift.  The  labor  cost  per  timber  is  thus  about 
$5  net,  or  $7  including  headmen,  engineers,  etc. 

Ring  Timber.  —  The  commonest  form  of  timber  lining, 
and  the  one  that  is  used  in  all  large  shafts,  consists  of  hori- 
zontal square-framed  sets,  spaced  4  to  6  ft.  centers  and 
lagged  with  2  to  3  in.  plank.  (See  Fig.  53.) 

In  a  previous  chapter,  the  terms  bunton,  end  plate,  wall 
plate,  and  punch  block  were  defined  as  the  cross  struts,  end 
timbers,  side  timbers,  and  posts,  respectively,  in  a  square- 
framed  set  of  timber.  Each  set  consists  of  two  wall  plates 


122  PRACTICAL  SHAFT  SINKING 

and  two  end  plates  halved  at  the  corners,  and  one  or 
more  buntons,  butted  against  the  inner  faces  of  the  wall 
plates  to  serve  as  struts.  The  buntons  also  divide  the  shafts 
into  the  requisite  number  of  compartments  and  afford  support 
for  guides,  etc.  The  sets  are  separated  by  posts  or  punch 
blocks  placed  at  the  corners  and  at  the  end  of  the  buntons. 

Lagging  plank  are  set  vertical  and  close  together.  They 
are  usually  placed  back  of  the  timber,  and  are  held  by  cord 
wood  or  slabs  packed  into  the  space  between  them  and  the 
rock.  This  packing  also  acts  as  a  support  for  the  sides  of 
the  excavation.  To  prevent  plank  from  falling  in  case  of 
displacement  of  the  packing,  they  should  be  well  spiked  to 
the  timbers  top  and  bottom.  The  best  construction  is 
effected  by  spiking  2  X  2  in.  strips  horizontally  in  the  middle 
of  the  outer  faces  of  the  wall  and  the  end  plates;  the  lagging 
boards  are  then  cut  two  inches  short  of  the  center  to  center 
spacing  of  the  sets,  and  are  inserted  between  the  strips. 

In  hard  rock  packing  is  not  necessary,  but  lagging  is 
usually  needed  to  prevent  water  from  splashing  into  the 
shaft  and  to  lead  it  to  the  rings.  In  this  case,  the  lagging 
boards  are  placed  between  the  timber  sets  and  are  held  by 
strips  nailed  to  the  top  and  bottom  faces  of  the  timbers. 
Sometimes  these  strips  are  beveled  so  as  to  lead  water  to 
the  back  side. 

When  a  shaft  has  been  sunk  without  support  as  far  as 
the  condition  of  the  sides  will  permit  —  say  50  to  100  ft.  — 
a  set  of  hitch  timbers  (dead  logs  or  bearing  timbers)  is  placed 
as  a  foundation,  and  the  timbering  is  built  up  through  the 
unlined  section.  The  hitch  timbers  are  set  into  hitches  cut 
in  the  rock  as  described  above,  and  must  be  perfectly  level 
and  in  line.  Since  they  are  to  carry  a  great  weight  the 
hitches  must  be  deep  enough  to  afford  a  bearing  equal  in 
strength  to  the  timber.  The  hitch  timbers  are  often  made 
deeper  than  the  ring  timbers;  8  X  12  in.  hitch  timbers,  for 
instance,  for  8  X  10  in.  ring  timbers.  A  heavy  floor  is 
built  over  the  hitch  timbers  between  the  first  set  of  timber 
and  the  wall,  to  carry  the  packing.  (See  Fig.  58.) 


SHAFT  LININGS 


123 


Each  set  of  timbers  must  be  securely  blocked  and  wedged 
in  the  corners  and  opposite  the  edge  of  the  buntons.  The 
lagging  and  packing  are  then  finished  before  the  timbers 
for  the  next  set  are  lowered.  The  workmen  stand  on  planks 
laid  across  the  buntons  and  raised  as  each  set  is  completed. 
In  order  to  avoid  splicing,  the  wall  plates  for  the  two  or 
three  closing  sets  (joining  to  completed  timbering  above) 
must  be  lowered  into  the  shaft,  laid  upon  the  timbers  already 
placed,  and  then  raised  horizontally. 


FIG.  55.  —  Brick  Shaft  Lining 

Corner  joints  should  be  framed  so  as  to  develop  the 
full  strength  of  the  timbers.  Punch  blocks  and  buntons  are 
usually  notched  into  the  wall  plates,  being  thus  secured 
laterally;  the  punch  blocks  are  allowed  to  extend  out  under 
the  ends  of  the  buntons  to  provide  vertical  support.  In 
addition,  keystoned  shape  notches  are  often  cut  in  the  wall 
plates,  into  which  the  ends  of  the  buntons  are  fitted. 

The  cost  of  placing  ring  timbers  of  this  type  varies  from 
$20  to  $30  per  M.  ft.  B.M. 

In  small  shafts,  the  lagging  plank  are  often  placed  hori- 
zontally and  spiked  together,  and  are  notched  into  each 
other  at  the  corners.  This  gives  all  the  support  that  is 
required  in  a  small  square  shaft  or  well;  in  a  compartment 


124  PRACTICAL  SHAFT  SINKING 

shaft,  pairs  of  vertical  timbers  serving  both  as  posts  and 
girders  are  placed  against  the  side,  and  buntons  are  set 
between  them.  Vertical  nailing  strips  are  also  put  in  the 
corners.  (See  Fig.  58a.) 

Brick.  —  The  thickness  of  a  brick  lining  varies  from 
9  to  18  in.,  depending  on  the  size  of  the  shaft  and  the  nature 
of  the  rock.  A  13-in.  wall  is  the  usual  thickness  for  a  16-ft. 
shaft.  The  wall  is  built  in  sections  50  to  100  ft.  long,  each 
of  which  is  founded  on  a  curb  of  wood  or  iron  built  in  seg- 
ments and  set  into  a  groove  in  the  side  of  the  shaft.  This 
groove  is  cut  by  hand,  the  bottom  being  made  perfectly 
level,  and  the  curb  is  carefully  wedged  into  shape  exactly 
concentric  with  the  shaft.  If  much  water  leaks  into  the 
shaft  in  the  section  that  is  to  be  lined,  the  curb  is  made  of 
iron  and  the  space  behind  it  is  filled  with  wooden  blocks 
and  wedges  driven  in  tight,  as  will  be  described  in  detail 
later.  A  water  ring  is  cast  on  the  inside  of  the  curb,  and 
drainage  pipes  leading  through  the  lining  to  the  ring  are 
provided.  (See  Fig.  55.) 

The  work  of  setting  the  curb  and  building  the  wall  is 
done  on  a  suspended  scaffold  which  is  raised  as  required  by  a 
special  engine  at  the  surface.  The  wall  is  built  very  rapidly 
as  several  masons  can  work  without  interference  —  six 
masons,  if  kept  supplied  with  brick  and  mortar,  can  easily 
build  6  to  8  ft.  of  wall  per  shift. 

In  deep  shafts  recently  sunk  in  England,  the  work  has 
been  so  arranged  that  sinking  and  lining  are  carried  on 
simultaneously.  Two  buckets  are  used,  one  of  which  hangs 
in  the  center  of  the  shaft,  and  passes  through  a  hole  in  the 
walling  scaffold.  The  other  hangs  sufficiently  off  center 
to  clear  the  sinking  bucket,  and  is  used  for  supplying  ma- 
terial to  the  masons.  The  ropes  which  suspend  the  scaffold 
also  serve  as  guides  for  the  buckets.  This  method  is  only 
feasible  for  large  shafts. 

The  brick  should  be  made  of  good  clay,  burned  hard 
enough  to  withstand  the  action  of  water.  Mortar  is  made 
of  quick-setting  Portland  or  hydraulic  cement  and  is  used 


SHAFT  LINING 


125 


as  sparingly  as  possible.     The  space  back  of  the  wall  is 
filled  with  concrete,  tamped  clay,  or  cinders. 

Iron   Tubbing.  —  Both    English    and    German    tubbing 
consists  of  cast-iron  segments  and  is  built  up  in  rings.     The 


Wedge 


i  - 

3      j 

jT}H 

I  u 

u           u 

FIG.  56.  —  English  Tubbing  Segment  —  15  to  Circumference 

English  segments,  however,  have  rough  edges  and  no  bolts 
and  are  made  water-tight  by  wedging  the  cracks  with  wood, 
whereas  the  German  have  machined  edges  provided  with 
flanges  and  are  bolted  together.  Lead  gaskets  are  used  to 
make  a  tight  joint.  English  segments  are  about  2  ft.  and 
German  about  5  ft.  high. 


-Bolt  Hole 


FIG.  57.  —  German  Tubbing  Segment  —  8  to  Circumference 

The  process  of  setting  a  wedging  curb  and  building  up 
English  tubbing  in  wet  rock  is  described  by  Mr.  J.  J.  Prest 

as  follows: 


126 


PRACTICAL  SHAFT  SINKING 


"The  shaft  was  sunk  about  9  or  10  ft.  into  the  imperme- 
able stratum,  of  the  full  diameter  of,  say,  23  ft.,  and  then 
decreased  abruptly  to  the  finished  size  of  the  shaft,  say 
20  ft.,  and  the  sinking  was  continued  a  further  distance  of 
6  or  8  ft.  The  cradle  (walling  scaffold)  was  then  lowered 


FIG.  58a.  —  Timber  Sketches 

into  the  pit  bottom  and  a  temporary  wood  water-ring  was 
fixed  on  dowels  about  9  or  10  ft.  above  the  site  selected  for 
the  bed  of  the  wedging  curb.  The  whole  of  the  water 
running  from  the  sides  of  the  shaft  was  then  collected  in  this 
temporary  water-ring  and  allowed  to  run  off  in  canvas  hogges 
or  trunks  at  two  or  three  different  positions  to  the  pumps. 


SHAFT  LINING  127 

"The  cradle  having  been  fixed  in  position,  the  sinkers 
proceeded  to  level  the  surface  of  the  rock  bed  with  mat- 
tocks, and  when  this  was  accomplished  to  the  satisfaction 
of  the  engineer,  a  wedging  curb,  three  segments  of  which 
were  fitted  with  valves  to  pass  gas  or  air  accumulated  behind 
the  tubbing,  was  laid  on  the  bare  rock,  seasoned  red-wood 
sheeting  f  in.  thick  was  placed  between  all  end  joints,  and 
the  spaces  between  the  back  of  the  curb  and  the  rock  were 
filled  with  dry-wood  gluts  to  bring  the  inside  of  the  curb  up 
to  the  finished  diameter  of  the  shaft.  Afterwards  well- 
seasoned  tapered  dry-wood  wedges  were  driven  into  the  wood 
packing  between  the  back  of  the  curb  and  the  strata  until 
steel  chisel  points  refused  to  enter.  Then  a  layer  of 
f-in.  horizontal  sheeting  was  placed  on  the  top  of  the  wedg- 
ing curb,  and  the  2-ft.  foundation  course  of  tubbing  was  put 
on,  breaking  joints  with  the  curb,  f-in.  red- wood  sheeting 
being  placed  between  the  end  and  horizontal  joints,  and  the 
course  was  brought  up  to  the  correct  radius  of  the  shaft  by 
means  of  wood  packing.  Next  one  or  two  courses  of  plain 
tubbing  were  put  on,  the  fourth  course  usually  containing 
three  or  more  special  segments  (technically  termed  'valve- 
segments'),  cast  with  holes  4  to  6  in.  in  diameter  in  the 
center,  with  the  object  of  permitting  the  water  to  pass  from 
the  back  to  the  front  side  of  the  tubbing,  and  so  to  the 
pumps,  when  the  temporary  wood  water-ring  was  removed. 

"The  next  operation  was  to  wedge  lightly  the  vertical 
joints  of  the  three  or  four  courses  of  tubbing,  and  to  run  the 
whole  up  solid  with  good  cement  grout.  The  temporary 
water-ring  was  then  removed,  and  additional  courses  of 
ordinary  tubbing  were  built  on,  to  a  total  height  of  about 
60  ft.  The  joints  were  now  lightly  wedged,  commencing 
from  the  top  with  the  vertical  joints,  and  from  the  bottom 
with  the  horizontal  seams,  using  red-wood  wedges.  Addi- 
tional courses  of  segments  of  suitable  height  being  used 
to  close  up  to  the  wedging  curb  above,  the  vertical  and 
horizontal  seams  were  again  twice  wedged  alternately  as 
before,  and  the  small  center  holes  in  each  segment  were 


128  PRACTICAL  SHAFT  SINKING 

plugged.  Finally  the  large  holes  in  the  valve  segments 
passing  the  feeders  were  plugged  simultaneously  with  long 
tapered  plugs  of  wood,  the  excess  being  sawn  off  flush  with 
the  front  side  of  the  orifice,  the  cast-iron  caps  were  bolted 
on  to  the  flanges  and  the  shafts  was  rendered  dry  if  the 
work  had  been  well  carried  out." 

German  tubbing  is  started  from  a  wedging  curb  and  is 
bolted  together  as  built  up.  When  the  rock  itself  is  treacher- 
ous as  well  as  wet,  under-hanging  tubbing  is  sometimes 
employed.  In  this  case  a  wedging  curb,  faced  on  the  bottom 
and  provided  with  bolt  holes,  is  set  just  above  the  shaft 
bottom.  The  segments  are  lowered  to  the  bottom,  and  are 
grasped  one  at  a  time  with  a  special  pair  of  tongs  and  raised 
to  the  under  side  of  the  curb.  Each  segment  is  then  sus- 
pended by  two  bolts  with  long  threads  and  the  tongs  are 
removed.  The  lead  gasket  is  inserted;  the  segment  is  raised 
to  place  by  screwing  up  the  nuts  on  the  long  bolts,  and  is 
then  bolted  up  tight.  The  process  is  repeated  as  soon  as 
the  shaft  has  been  deepened  sufficiently.  After  each  ring 
has  been  bolted  up,  the  opening  at  the  bottom  between  it 
and  the  rock  is  closed  with  plates  and  wedges,  and  the 
space  behind  is  filled  with  cement  grout  poured  in  through 
holes  in  the  segments. 


CHAPTER  X 

CONCKETE  LININGS.    COSTS  PER  LINEAR  FOOT  FOR  RECT- 
ANGULAR, ELLIPTICAL,  AND  QUADRILATERAL  SHAFTS 

Two  types  of  concreted  shafts  are  to  be  considered:  the 
circular  or  elliptical,  with  unsupported  lining;  and  the 
rectangular,  with  reinforced  concrete  lining  supported  by 
steel  beams,  concrete  buntons,  or  walls.  In  both,  the  con- 
crete is  placed  directly  against  the  rock  walls  and  an  inner 
form  only  is  required.  From  a  construction  standpoint 
the  types  are  equally  feasible,  and  the  choice  depends  upon 
the  cost.  In  several  cases  known  to  the  writer  a  com- 
promise has  been  effected  by  shaping  the  shaft  as  a  quadri- 
lateral with  sides  formed  of  circular  arcs. 

For  a  single  compartment  air  shaft  the  circular  shape 
is  in  every  way  the  most  desirable,  not  only  because  the 
circular  shaft  is  cheaper  to  sink  than  a  square  shaft  of  equal 
area,  but  also  because  a  circular  ring  of  plain  concrete  is  the 
strongest  lining  possible  with  a  given  amount  of  material. 

In  the  case  of  a  shaft  with  two  or  more  compartments, 
the  selection  of  the  most  economical  shape  requires  some 
calculation.  At  first  sight  it  would  seem  that  a  simple 
rectangular  shaft,  surrounded  by  a  concrete  wall  only  thick 
enough  to  be  as  strong  as  the  usual  timber  lining,  would  be 
a  satisfactory,  as  well  as  a  cheap,  shape,  but  this  is  not  the 
case.  A  concrete  lining,  even  when  provided  with  weep 
holes,  must  resist  some  hydrostatic  pressure;  a  timber 
lining  has  none  to  resist.  Furthermore,  permanent  weep 
holes  are  most  undesirable;  the  concrete  should  exclude  the 
water  entirely,  and  hence  must  be  designed  to  bear  very 
great  pressure  at  considerable  depth.  Just  what  amount 
the  theoretical  pressure  is  reduced  by  the  adhesion  of  the 

129 


130 


PRACTICAL  SHAFT  SINKING 


concrete  to  the  shaft  walls  and  by  the  blocking  of  the  fissures 
with  grout  cannot  be  calculated.     In  solid  rock,  where  the 


FIG.  59.  —  Rectangular  Concrete  Lined  Shaft 

water  enters  in  well-defined  springs,  the  proper  grouting 
of  the  springs  will  relieve  the  lining  of  all  pressure.     In  very 

TABLE   1.     QUANTITIES  AND  COSTS  OF  RECTANGULAR  SHAFT 


Depth  in  feet  

20 

50 

100 

150 

200 

Total  thickness  of  lining  in  inches 

14 

21 

28 

34 

39 

Quantities  per  linear  foot: 

Concrete  to  neat  line  in  cu.  yds. 

3.90 

5.70 

7.60 

9.30 

10.70 

Concrete,  actual  in  cu.  yds  

5.80 

7.70 

9.70 

11.50 

13.00 

Excavation  to  neat  line  in  cu. 

. 

yds  

12.80 

14.60 

16.50 

18.20 

19.70 

Excavation,  actual  in  cu.  yds.  .  .  . 

14.70 

16.60 

18.60 

20.40 

22:00 

Weight      reinforcing      steel      in 

pounds  

256 

443 

650 

845 

1,030 

Costs  per  linear  foot: 

Forms    

$  25.00 

$  25.00 

$  25.00 

$  25.00 

$  25.00 

Concrete  at  $5  cu.  yd  

29.00 

38.50 

48.50 

57.50 

65.00 

Excavation  (see  note  *)  

49.60 

53.20 

57.00 

60.40 

63.40 

Reinforcing  steel  at  $.02  Ib  

5.10 

8.90 

13.0 

16.90 

20.60 

Total  

$108.70 

$125.60 

$143.50 

$159.80 

$174.00 

seamy  rock,  on  the  other  hand,  the  lining  may  have  to  bear 
parctically  the  full  hydrostatic  pressure. 

In  order  to  compare  the  costs  of  the  different  shapes,  let 

*  NOTE.  —  Cost  of  excavation  figured  on  basis  of  $4  per  cubic  yard  for 
section  containing  12  yards  per  linear  foot;  additional  excavation  at  $2  per 
cubic  yard.  Thus,  cost  of  16  cubic-yard  section  =  12  X  $4  +  4  X  $2  =  $56. 
These  unit  costs  are  for  purposes  of  comparison  only  and  should  not  be 
used  for  estimating. 


CONCRETE  LININGS 


131 


us  consider  in  detail  three  designs  for  a  shaft  with  two 
7  X  10  ft.  hoist  ways  and  an  airway  with  an  area  of  100 


FIG.  60.  —  Elliptical  Concrete  Lined  Shaft 

sq.  ft.     As  the  whole  area  of  a  hoist  shaft  is  ordinarily 
used  for  the  passage  of  air,  the  size  of  the  air  compartment 

TABLE  2.     QUANTITIES  AND  COSTS  OF  ELLIPTICAL  SHAFT 


(  0  to 

Depth  in  feet,  

(  100 

150 

200 

250 

300 

400 

Thickness  of  lining  in  inches, 

ends  

12 

12 

12 

12 

12 

12 

Thickness  of  lining  in  inches, 

sides    

12 

18 

24 

29 

34 

42 

Quantities  per  linear  foot: 

Concrete  to  neat  line,  cu. 

yds  

2.60 

3.40 

4.30 

5.00 

5.70 

6.80 

Concrete,  actual  in  cu.  yds. 

4.40 

5.20 

6.10 

6.80 

7.50 

8.60 

Excavation  to  neat  line  in 

cu.  yds  

15.20 

16.00 

16.90 

17.60 

18.30 

19.40 

Excavation,  actual  in   cu. 

yds  

17.00 

17.80 

18.70 

19.40 

20.10 

21.20 

Costs  per  linear  foot  : 

Forms  

$15.00 

$15.00 

$15.00 

$15.00 

$15.00 

$15.00 

Concrete  at  $5  cu.  yd.    ... 

22.00 

26.00 

30.50 

34.00 

37.50 

43.00 

Excavation  (see  note*)  .  .  . 

54.40 

56.00 

57.80 

59.20 

60.60 

••  62.80 

Total  

$91.40 

$97.00 

$103.30 

$108.20 

$113.10 

$120.80 

may  be  reduced  if  the  rest  of  the  shaft  is  enlarged;  the  air- 
way must,  however,  be  large  enough  to  contain  pipes  and 


132 


PRACTICAL  SHAFT  SINKING 


ladders  and  to  provide  in  addition  an  ample  passage  for  air 
if  the  hoistways  are  temporarily  closed. 


FIG.  61.  —  Quadrilateral  Shaft 

Let  us  assume  a  minimum  thickness  of  12  in.  of  concrete 
for  a  water-tight  lining;  also  that  in  each  case  the  lining 

TABLE  3.     QUANTITIES  AND  COSTS  OF  QUADRILATERAL  SHAFT 


Depth  in  feet  
Thickness     of     lining     in 

fOto 
1  100 

150 

200 

250 

300 

400 

inches  

12 

19 

26 

32 

39 

52 

Quantities  per  linear  foot  : 

Concrete  to  neat  line  in 

cu.  yds  

2.70 

4.40 

6.20 

7.90 

9.90 

13.90 

Concrete,    actual   in    cu. 

yds  

4.50 

6.30 

8.20 

10.00 

12.10 

16.20 

Excavation  to  neat  line  in 

cu.  yds  

14.90 

16.60 

18.40 

20.10 

22.10 

26.10 

Excavation,  actual  in  cu. 

yds  

16.70 

18.50 

20.40 

22.20 

24.30 

28.40 

Costs  per  linear  foot: 

Forms  

$15.00 

$15.00 

$15.00 

$15.00 

$15.00 

$15.00 

Concrete  at  $5  cu.  yd.    . 

22.50 

31.50 

41.00 

50.00 

60.50 

81.00 

Excavation  (see  note*)  .  . 

53.80 

57.20 

60.80 

64.20 

68.20 

76.20 

Total 

$91.30 

$103.70 

$116.80 

$129.20  $143.70 

$172.20 

carries  the  entire  hydrostatic  pressure;  then  the  specifica- 
tions for  the  three  forms  of  shafts  will  be  as  follows: 

Rectangular  Shaft.  —  Fig.  59.  Two  hoistways,  7  X  10  ft.; 


CONCRETE  LININGS  133 

one  airway,  10  X  10  ft.  Ten-inch  concrete  dividing  walls 
in  place  of  buntons.  Extreme  inside  dimensions,  10  X  25  ft. 
8  in.  Area  airway,  100-sq.  ft.;  total  clear  area,  240  sq.  ft. 
Thickness  of  lining  at  any  point  made  equal  to  depth  of 
simple  beam  of  10-ft.  span  required  to  sustain  hydrostatic 
pressure  at  that  point.  Resisting  moment  and  weight  of 
reinforcement  calculated  by  Johnson's  formula,  factor  of 
safety  3.  (Ultimate  tensile  strength  of  steel,  65,000  Ibs. 
per  square  inch,  compressive  strength  of  concrete  in  beam, 
2500  per  square  inch.)  Reinforcing  steel  set  3  in.  inside 
of  face  of  wall. 

Cost  of  forms,  Table  1,  includes  cost  of  forms  for  dividing 
walls,  and  is  therefore  greater  than  the  cost  in  the  elliptical 
shafts. 

Excess  of  actual  over  theoretical  quantity  of  excavation 
is  estimated  as  15  per  cent,  for  28-ft.  shaft.  This  excess 
increases  with  the  length  of  the  shaft  only,  as  the  ends  are 
drilled  to  line. 

Elliptical  Shaft.  —  Fig.  60.  Extreme  inside  dimensions, 
16  X  27  ft.  Area  of  airway,  78  sq.  ft.  Total  clear  area, 
allowing  for  10-in.  buntons,  304  sq.  ft. 

Strength  of  lining  calculated  on  the  assumption  that  the 
stress  in  the  elliptical  cylinder  at  any  point  is  equal  to  that 
caused  in  a  circular  cylinder,  with  a  radius  equal  to  the 
radius  of  curvature  of  the  ellipse  at  the  given  point,  by  the 
same  hydrostatic  pressure  acting  upon  it.  The  lining  is 
therefore  made  thicker  at  the  sides  than  at  the  ends. 

To  prove  this  proposition,  assume  the  lining  to  be  con- 
structed of  a  number  of  small  portions,  each  the  arc  of  a 
circle.  The  stress  in  each  portion  caused  by  the  hydro- 
static pressure  of  the  film  of  water  between  it  and  the  rock 
is  directly  proportional  to  the  radius,  and  the  thickness  of 
each  section  should  therefore  be  made  proportional  to  the 
radius.  Considering  any  portion,  as  a-b,  Fig.  62,  the  skew- 
back  toward  the  side  of  the  ellipse  is  formed  entirely  by  the 
adjoining  portion,  while  the  skewback  toward  the  end  is 
formed  partly  by  the  adjoining  portion  and  partly  by  the 


134 


PRACTICAL  SHAFT  SINKING 


rock.  If  the  number  of  circular  portions  is  indefinitely 
increased,  the  unbalanced  end  thrust  of  each  will  be  taken 
up  by  the  irregularities  of  the  rock. 

Ultimate  compressive  strength  of  concrete,  3000  Ibs. 
per  square  inch;  factor  of  safety,  3. 

Excess  of  actual  over  theoretical  excavation  assumed  as 
12  per  cent,  for  smallest  section.  As  the  length  of  the  shaft 
does  not  vary,  this  excess  is  constant. 

Quadrilateral  Shaft.  —  Fig.  55.  Inside  dimensions,  16 
X  24  ft.  8  in.  Radius  of  ends  and  sides,  23  ft.  Area  of 
airway,  94  sq.  ft.  Total  clear  area,  allowing  for  10-in. 
buntons,  294  sq.  ft. 


FIG.  62 

For  calculating  stresses,  sides  and  ends  are  considered 
as  portions  of  a  46-ft.  circular  cylinder.  Ultimate  compres- 
sive strength  of  concrete,  3000  Ibs.  per  square  inch ;  factor  of 
safety,  3. 

Excess  of  actual  over  theoretical  quantity  of  excavation 
assumed  to  be  12  per  cent,  for  minimum  length  and  to 
increase  with  the  length. 

Methods  of  Working.  —  The  easiest  way  to  concrete  a 
shaft  is  to  finish  sinking,  then  start  at  the  bottom  and  build 
up.  Unfortunately  this  is  feasible  only  for  comparatively 
shallow  shafts  in  hard  dry  rocks,  and,  ordinarily,  successive 
lengths  of  lining  must  be  placed  as  the  shaft  deepens,  to 
protect  the  sides  and  cut  off  feeders  of  water. 


CONCRETE  LININGS  135 

When  the  shaft  has  been  sunk  as  far  as  seems  safe  or 
desirable,  hitches  are  cut  in  the  sides,  a  set  of  bearing  timbers 
is  placed  and  a  heavy  plank  floor  laid  upon  them  to  support 
the  concrete.  In  order  to  avoid  injury  to  the  concrete 
when  sinking  is  resumed,  the  platform  should  be  built  15  or 
20  ft.  above  the  shaft  bottom.  It  is  cheapest  to  cut  the 
hitches  near  the  bottom  when  the  shaft  is  at  the  proper 
depth  and  then  to  sink  three  or  four  cuts  further;  scaffolding 
is  thus  made  unnecessary  as  it  is  easy  to  place  timbers  in 
hitches  already  prepared.  In  a  rectangular  or  elliptical 
shaft  transverse  bearing  timbers  are  placed;  in  a  circular 
shaft  four  timbers  placed  in  the  form  of  a  square  make  the 
best  platform.  In  any  case  an  opening  should  be  left  in 
the  bucket  way  so  that  the  bottom  is  accessible. 

Forms  are  started  from  the  platform  and  are  built  in 
rings  5  to  10  ft.  high.  When  each  ring  of  forms  is  completed 
a  temporary  floor  is  laid  on  top  of  it.  Concrete  mixed  at 
the  top  of  the  shaft  is  lowered  in  shaft  buckets  and  dumped 
on  the  floor,  whence  it  is  shoveled  behind  the  forms.  To 
prevent  loss  of  concrete  the  floor  must  be  laid  with  tight 
joints,  and  this  is  most  easily  accomplished  by  making  it 
in  sections  as  large  as  can  be  lowered  into  the  shaft.  An- 
other plan  is  to  dump  the  concrete  from  the  bucket  into  the 
forms  direct  through  a  movable  chute;  the  necessity  for 
laying  the  tight  floor  is  thus  obviated. 

A  |-yd.  batch  mixer  (such  as  the  Smith  or  Ransome) 
gives  the  best  results.  One-half  yard  of  wet  concrete  is 
about  all  an  ordinary  shaft  bucket  will  hold  without  spilling. 
The  use  of  a  batch  mixer  makes  it  possible  to  use  only  one 
bucket  without  losing  time,  as  a  batch  is  being  lowered  and 
dumped  while  the  next  is  being  mixed.  The  mixer  should 
if  possible  be  set  below  the  ground  level  (see  Fig.  63)  to 
avoid  elevating  the  materials.  Below  the  mixer  a  plat- 
form is  placed  for  supporting  the  bucket  while  it  is  being 
filled.  This  platform  should  be  3  ft.  wide  and  so  situated 
that  a  bucket  hanging  free  on  the  rope  will  clear  it  10  or 
12  in.,  and  should  be  provided  with  a  hand  rail  except  for 


136 


PRACTICAL  SHAFT   SINKING 


4  ft.  immediately  in  front  of  the  mixer.  When  concreting, 
two  men  stand  on  the  platform,  one  on  either  side ;  the  empty 
bucket  is  hoisted  slightly  above  the  platform  level,  is  grasped 


FIG.  63.  —  Lowering  and  Placing  Concrete 

by  the  two  men  and  swung  in  under  the  discharge  chute  of 
the  mixer,  and  is  then  lowered.  When  filled  it  is  hoisted 
slightly,  swings  out  over  the  shaft  and  is  steadied  by  the  men 
on  the  platform  before  being  lowered. 


CONCRETE  LININGS  137 

By  working  as  outlined  above  with  a  good  organization 
it  is  easy  to  place  12  or  15  one-half  yard  batches  per  hour, 
and  a  shaft  lining  containing  as  much  as  6  yds.  of  concrete 
per  linear  foot  can  thus  be  placed  at  the  rate  of  more  than  a 
foot  an  hour  —  about  as  fast  as  timber.  It  is  the  time 
required  to  construct  the  foundation  platform,  to  set  the 
forms,  and  to  connect  a  new  section  of  lining  to  the  one 
above  it  that  makes  the  process  slow. 

Foundation  platforms  must  be  built  and  closures  must 
be  made;  time  can  be  saved  only  by  reducing  the  number. 
Plenty  of  forms  should  therefore  be  provided,  and  the  sec- 
tions made  as  long  as  possible.  The  design  of  forms,  both 
as  regards  strength  and  finish  and  facility  of  erection, 
demands  careful  consideration. 

Forms.  —  The  writer  has  had  experience  with  three 
types  of  forms.  The  first,  consisting  of  wooden  slabs,  was 
used  for  lining  a  17  X  33  ft.  elliptical  shaft  at  Tug  River, 
W.  Va.  (Engineering  News,  November  7,  1904).  The 
slabs  were  made  of  2-in.  vertical  lagging  planks  nailed  top 
and  bottom  to  double  2  X  12  in.  centers.  They  were  5  ft. 
high,  and  8  slabs  made  up  a  complete  ring.  These  forms 
made  a  satisfactory  wall,  but  were  heavy  and  were  greatly 
damaged  by  moving.  Eight  to  ten  hours  were  required  to 
set  one  ring;  the  work  was  divided  into  two  shifts,  one 
setting  forms  and  one  concreting,  so  the  progress  made  was 
only  5  ft.  per  day. 

The  second  type,  consisting  of  steel  slabs,  was  used  on 
several  waterway  shafts  on  the  Catskill  Aqueduct.  These 
shafts  are  circular,  about  14  ft.  6  in.  in  finished  diameter, 
and  a  perfectly  smooth  surface  is  required.  The  slabs  are 
5  ft.  high  and  are  made  four  to  a  ring.  Forty  feet  of  forms 
were  provided  for  each  shaft.  Two  rings  were  erected  at  a 
time,  the  concrete  being  tamped  and  spaded  with  long- 
handled  tools,  and  when  new  it  was  possible  to  erect  and 
fill  two  rings  in  twelve  hours.  With  use  the  forms  become 
more  difficult  to  set:  the  progress,  including  platforms  and 
closures,  averaged  10  ft.  per  day.  A  pair  of  wooden  key 


138  PRACTICAL  SHAFT  SINKING 

blocks  is  provided  at  opposite  joints  in  each  ring.  These 
are  chopped  out  to  release  the  forms. 

After  a  section  of  lining  is  finished,  the  steel  slabs  are 
usually  left  in  place  until  the  next  section  is  ready  and  then 
moved  down,  a  ring  at  a  time.  If  the  section  to  be  concreted 
is  longer  than  the  available  forms,  a  working  platform  may 
be  suspended  below  the  forms  with  ropes  and  the  slabs 
taken  off  at  the  bottom  and  moved  up.  The  slabs  should 
always  be  cleaned  and  oiled  before  being  used  again. 

The  third  type  consists  of  angle-iron  rings  spaced  4-ft. 
centers  and  lagged  with  vertical  2-in.  plank.  Wooden 
nailing  strips  are  bolted  to  the  angles.  These  forms, 
although  they  make  a  surface  inferior  to  the  steel  forms,  are 
cheaper  and  easier  to  place.  One  ring  at  a  time  is  set  and 
filled,  and  by  working  continuously  four  or  five  can  be 
completed  in  twenty-four  hours.  These  forms  are  removed, 
taken  to  the  surface  and  cleaned  after  each  section  of  lining 
is  finished. 

Placing.  —  Concrete  should  be  mixed  wet,  even  though 
considerable  water  is  present  in  the  shaft,  and  should  be 
thoroughly  spaded  next  the  forms.  Springs  of  any  volume 
appearing  in  a  section  that  is  to  be  lined  should  be  taken 
care  of  in  advance  of  the  lining :  this  can  be  done  by  drilling 
a  hole  into  the  water-bearing  seam,  inserting  a  pipe  long 
enough  to  carry  the  water  out  into  the  shaft,  and  caulking 
around  the  pipe.  All  springs  should  be  piped  through  the 
forms  into  the  shaft  before  concrete  is  placed,  Fig.  63a, 
for  if  this  precaution  is  neglected  a  very  slight  inflow  may 
accumulate  enough  head  to  disrupt  the  green  concrete  and 
cave  in  the  forms.  The  writer  has  known  this  to  actually 
happen  and  has  had  to  dig  out  about  80  sq.  ft.  of  lining  dis- 
placed in  this  way  by  the  water  from  one  tiny  spring. 

Grouting.* — The  drainage  pipes  should  be  provided  at 
the  inner  ends  with  sleeves  set  flush  with  the  face  of  the  wall. 
If  the  concrete  has  been  properly  placed  and  too  much 
water  has  not  been  encountered,  the  shaft  can  be  made  dry 
by  plugging  the  sleeves.  If,  however,  the  concrete  is  .porous, 

*  See  Appendix  A. 


CONCRETE  LININGS  139 

it  may  be  necessary  to  force  grout  through  the  pipes  until 
all  crevices  are  filled.  The  grout  is  mixed  very  thin,  and 
can  be  pumped  in  with  a  high-pressure  pump  or  expelled 
from  a  tank  and  driven  into  the  pipe  by  the  use  of  high-pres- 
sure compressed  air.  Concrete  can  be  made  absolutely 
water-tight  in  this  way  if  enough  grout  pipes  are  provided. 
In  closing,  the  writer  wishes  to  express  his  gratitude  to 
Mr.  Evan  Edwards,  of  Scranton,  whose  knowledge  of  prac- 
tical shaft  sinking  has  been  of  the  greatest  assistance  in  the 
preparation  of  this  book. 


APPENDIX  A 

GROUTING  SHAFTS  4  AND  24,  NEW  YORK 
CITY  AQUEDUCT 

SHAFT  4 

SHAFT  4  is  concrete-lined  circular  shaft  sunk  through  the 
Fordham  Gneiss  near  Jerome  Park  Reservoir.  The  depth 
to  tunnel  invert  is  242  ft.;  the  finished  diameter  of  the 
concrete  lining  is  14'  0",  and  the  effective  average  thick- 
ness of  the  lining  in  normal  rock  is  13".  Sinking  progressed 
without  incident  on  the  "one  drilling  and  two  mucking 
shift"  plan  until  a  depth  of  149  feet  was  reached;  in  the 
first  round  drilled  below  this  depth  clear  water  was  found 
in  four  of  the  thirty  holes.  The  rock  was  solid  and  each  hole 
was  plugged  with  a  wooden  plug  wrapped  in  sacking  as 
soon  as  the  bottom  was  reached. 

It  was  thought  that  the  water  might  be  only  a  pocket, 
so  after  a  proper  pump  had  been  placed  one  of  the  plugs 
was  removed.  Pumping  was  continued  for  two  weeks, 
during  which  time  the  upper  part  of  the  shaft  was  lined  and 
appliances  for  grouting  obtained.  As  the  flow  was  not 
appreciably  diminished,  it  was  then  decided  to  grout  the 
water-bearing  crevice  through  the  drill  holes. 

The  grout  mixer,  which  was  of  the  air  stirring  type 
standard  on  the  aqueduct,  was  lowered  to  the  shaft  bottom 
and  coupled  to  the  air  line,  and  was  then  connected  succes- 
sively to  the  various  wet  holes.  In  order  to  make  connec- 
tions, pieces  of  2-inch  pipe  3  feet  long  with  standard  threads 
at  one  end  were  given  a  gradual  taper  at  the  other;  as 
soon  as  a  plug  was  drawn  from  a  hole  the  tapered  end  of  the 
pipe  was  wrapped  with  sacking  and  driven  in  tight,  and  a 
2-inch  stop  cock  screwed  on.  The  mixer  was  attached  to 
the  holes  by  a  heavy  hose.  The  first  series  of  holes  was 

140 


APPENDIX   A  141 

grouted  in  one  shift;  the  next  day  more  holes  were  drilled 
and  more  clear  water  met  with,  and  grouting  was  resumed 
that  night.  After  12  hours  the  holes  refused  to  take  any 
more  grout,  and  operations  were  discontinued  for  24  hours. 
Regular  sinking  was  then  resumed  and  no  more  water  was 
met  for  a  week.  About  200  sacks  of  neat  cement  were  used 
in  blocking  this  fissure. 

At  a  depth  of  181  feet  the  water  was  again  found  in  the 
sump  holes,  and  grouting  was  once  more  necessary.  This 
time  the  water  was  heavily  charged  with  sand  and  pieces  of 
disintegrated  rock;  the  rock  around  the  collars  of  the  drill 
holes  was  seamy  and  it  was  harder  both  to  make  good 
connections  and  to  force  grout  into  the  holes  after  the  con- 
nectiong  were  made.  Grouting  was  nevertheless  persisted 
in  and  in  three  weeks  130  holes  from  10  to  20  feet  deep  were 
drilled  in  the  shaft  bottom,  and  over  50  cubic  yards  of  grout 
(mostly  neat  cement)  was  forced  into  the  fissures  at  pres- 
sures up  to  240  pounds  per  square  inch.  Almost  every 
hole  drilled  met  some  water,  and  the  usual  procedure  was 
to  drill  five  or  six  holes,  grout  them,  lay  off  a  shift  or  two 
to  permit  the  grout  to  set,  and  then  to  drill  five  or  six  more 
holes.  The  depth  at  which  water  was  met  gradually  in- 
creased, and  it  was  therefore  possible  to  sink  four  feet 
during  the  three  weeks  of  grouting.  The  leakage  over  the 
shaft  bottom,  however,  gradually  increased  to  85  gallons 
a  minute. 

At  the  end  of  this  period  the  shaft  bottom  was  so 
thoroughly  perforated  that ''there  was  no  sound  rock  left 
to  drill  into  and,  after  an  unsuccessful  attempt  to  stop  the 
leakage  with  a  concrete  blanket,  sinking  was  resumed.  It 
was  found  in  passing  through  that  the  disintegrated  area 
was  a  band  varying  from  1  to  5  feet  in  width,  circling  the 
shaft,  10  feet  higher  on  the  east  than  on  the  west  side  and 
twisted  and  folded  in  every  direction.  Below  this  was 
solid  and  perfectly  sound  rock. 

The  sand  was  compacted  so  thoroughly  by  the  pressure 


142  APPENDIX   A 

to  which  it  had  been  subjected  that  it  showed  no  tendency 
to  wash  out,  and  the  flow  into  the  shaft  did  not  appreciably 
increase. 

After  excavating  10  feet  below  into  the  solid  rock,  the 
shaft  was  concreted;  a  special  section  with  a  minimum  of 
24  inches  of  concrete  was  placed  for  20  feet,  extending  into 
the  solid  rock  above  and  below.  Reinforcing  steel,  con- 
sisting of  1-inch  rods,  24  inches  apart,  vertically  and  hori- 
zontally, was  placed  18  inches  from  the  face  to  prevent 
cracking.  The  disintegrated  area  was  first  lined  with  thin 
sheet  iron  and  the  water  was  led  through  the  forms  before 
concreting.  The  space  back  of  this  sheet  iron  was  packed 
with  rock.  Five  days  after  concreting  the  holes  were  plugged, 
and  the  gauge  showed  77  pounds  per  square  inch  pressure 
back  of  the  lining.  The  leakage  through  the  concrete  after 
the  holes  were  plugged  was  only  one  or  two  gallons  per 
minute.  When  the  concrete  had  set  for  a  month  this  leakage 
was  entirely  cut  off  by  grouting  the  holes. 

SHAFT  24 

Shaft  24  is  of  the  same  general  construction  as  Shaft  4 
and  penetrates  blocky  and  seamy  granodiorite  for  its  entire 
depth  in  the  rock,  Elevation-76  to  Elevation-269.  At 
intervals  horizontal  seams  were  encountered  generally 
\  inch  to  |  inch  thick  which  were  filled  with  finely  crushed 
rock  that  allowed  water  to  flow  into  the  shaft,  usually  in 
only  one  point  of  the  seam.  As  the  seams  were  encountered 
holes  were  drilled  into  them  2  or  3  feet  deep  wherever  there 
was  any  leakage  and  when  the  shaft  was  concreted  these  holes 
were  fitted  with  grout  pipes  and  were  grouted.  No  great 
quantity  of  water  was  met  with  until  at  Elevation-221, 
while  drilling  a  round,  a  flow  of  240  gallons  a  minute  was 
struck,  all  the  water  coming  out  of  one  hole.  This  water 
was  under  the  full  hydrostatic  pressure  due  to  the  head  and 
flooded  the  shaft.  After  about  a  week's  delay,  in  which 


APPENDIX    A  143 

time  large  pumps  were  obtained,  the  shaft  was  emptied. 
Grouting  was  then  commenced,  using  the  same  general 
methods  as  were  used  at  Shaft  4.  In  all,  10  holes  were  drilled 
and  75  cubic  yards  of  neat  cement  grouting  was  pumped 
in  under  pressures  up  to  400  pounds  per  square  inch.  The 
leakage  into  the  shaft  was  reduced  by  this  grouting  to 
about  25  gallons  per  minute. 


APPENDIX  B 

SEVERAL  of  the  shafts  of  the  City  Aqueduct  on  the  lower 
east  side  of  Manhattan  Island  and  in  Brooklyn  were  sunk 
through  great  depths  of  water-bearing  sand  by  the  pneu- 
matic caisson  method.  The  caissons  are  of  two  sizes,  15  feet 
4  inches  and  18  feet  in  diameter,  and  2  feet  and  3  feet  thick 
respectively,  and  are  heavily  reinforced  with  vertical  and 
horizontal  rods.  At  the  bottom  of  each  a  heavy  steel  shoe 
is  imbedded  to  form  a  cutting  edge.  The  vertical  reinforc- 
ing rods  are  attached  to  the  shoe.  The  roof  of  the  working 
chamber  is  formed  by  a  heavy  reinforced  concrete  deck 
with  two  openings,  to  which  the  air  shafts  for  the  man  lock 
and  for  the  material  lock  are  connected. 

Each  shaft  chamber  was  first  excavated  and  timbered 
square  to  about  the  ground  water  level,  and  the  caisson  was 
then  started  from  the  bottom  of  the  chamber  and  built  up 
to  the  full  height  before  sinking.  Steel  forms  were  used 
throughout.  The  deepest  caisson  projected  over  60  feet 
above  the  ground  before  sinking  was  started. 

Excavated  material  was  handled  by  a  stiff-leg  derrick. 
The  caisson  was  loaded  with  pig  iron,  and  the  space  between 
the  shafting  and  the  concrete  filled  with  excavated  sand 
for  additional  weight.  Sinking  proceeded  rapidly  in  three 
shifts,  at  one  of  the  shafts  rock  being  reached  through 
55  feet  of  sand  in  five  days. 

The  most  difficult  part  of  the  operation  is  the  sealing 
of  the  caisson  into  the  rock,  so  that  the  air  can  be  safely 
taken  off.  To  accomplish  this  the  ledge  is  leveled  off  and 
the  rock  excavated  for  several  feet,  the  caisson  meanwhile 
being  supported  on  timber  posts.  The  rock  is  then  thor- 
oughly cleaned  with  high  pressure  air,  a  one-to-two  mortar 
collar  is  built  around  the  shaft  |  of  an  inch  back  of  the 

144 


APPENDIX    B 


145 


outer  edge  of  the  shoe,  and  the  posts  are  then  shot  out, 
allowing  the  caisson  to  sink  through  the  collar  into  its  final 


CITY  TUNNEL 

Compressed  Air  Caiss. 

Appendix  B 


position.    The  space  between  the  collar  and  the  caisson  is 
then  grouted  through  pipes  previously  set  in  the  collar. 


INDEX 


After-cooler,  27. 

Air-box  ventilator,  79. 

Air  compressor,  Corliss,  27. 
straight-line,  26. 
two-stage,  26. 

Air-drills,  2,  72-75. 
hose,  78. 
lift,  88. 

-     lock,  60. 

Anhalt  Government  Salt  Mine,  freez- 
ing, 93. 

Anzin  colliery  cementation,  105. 

Bearing  timbers,  41. 

Bell,  24. 

Bench,  67. 

Bethune  colliery  cementation,  105. 

Billy,  25. 

Bits,  drill,  70,  73. 

Blacksmith-shop  tools,  31. 

Blasting,  66,  68,  69. 

Blowers,  29,  80. 

Blow  pipe,  62. 

Boilers,  17. 

Bore  holes,  9,  92-94. 

Borers,  percussion,  88. 

sack,  86. 
Boring  process,  97-102. 

progress,  102. 

tools,  100,  101. 
Brakes,  23. 
Brick  lining,  124. 
Buckets,  21,  22. 
Buildings,  29. 
Bull  chain,  21. 
Buntons,  35,  119,  121. 

Caissons,  43-63 
circular,  44. 
open,  43-58. 


Caissons,  pneumatic,  59-63. 

rectangular,  43. 
Calumet  shaft,  4. 
Carnegie  steel  sheet  piling,  39. 
Central  power  plant,  27. 
Cementation,  102-106. 
Chambers,  W.  H.,  112. 
Chapin  Mine  Company  freezing,  93 
Chaudron,  M.  J.,  97. 
Churn  drill,  71. 
Chute,  21. 

Circular  concrete  lined  shaft,  129. 
Column,  75. 

arm,  75. 

Compound  sinking  drum,  90. 
Compressors,  air,  26. 
Concrete,  43,  129-139. 

forms,  137,  138. 

mixer,  135. 

placing,  138. 

shaft  linings,  129-139. 

timbers,  119. 
Conisboro  sinking,  112. 
Contract,  features  of,  5,  6. 

form  of,  10-12. 
Contract  prices,  84. 
Corliss  air  compressor,  27. 
Cost  of  concrete  lining,  130-132. 

plant,  32. 

sinking  in  rock,  81-83. 

sinking  through  surface,  49-51. 

timber,  121,  123. 
Curb  for  brick  lining,  124. 

D.  L.  &  W.  caisson,  43,  54-58. 
Derrick,  20. 

Direct  cementation,  102-106. 
Drainage  of  rock  water,  138. 

surface  water,  34,  42,  43. 
Drill,  air,  2,  72-75.  ' 


147 


148 


INDEX 


Drill,  diamond,  9. 

hand,  70,  71. 
Drill   holes,   depth   and   inclination, 

68,  69. 
Drill  mountings,  75,  76. 

steel,  70,  73. 
Drilling  frame,  76. 
Dummy,  25. 
Dump  car,  21,  22. 
Dynamite,  2,  67. 

thawing,  30. 

Electric  hoist,  4,  24. 

light,  27,  29. 

Elliptical  shaft,  131,  133. 
End  plate,  35,  121. 
Engines,  hoist,  20,  23,  24. 
Excavating  lock,  62. 
Excavation  under  air,  60. 
Explosives,  2. 

Fan,  29,  80. 

Feed-pump,  18. 

Feedwater  heater,  18. 

Fire  and  water  excavating,  1. 

Firm  earth,  35,  36. 

Fore  poling,  41. 

Forms,  concrete,  137,  138. 

Freezing  period,  93. 

Freezing  process,  35,  91-96. 

Gas  engine,  24. 
Gelatin,  67. 
Generator,  27. 
Grab  bucket,  86. 
Grout,  105,  138. 

mixer,  105. 

tank,  106. 
Guides,  25. 
Gunpowder,  2. 

Hand  drill,  70,71. 
Hanging  bolts,  36. 
Haniel  &  Lueg,  97. 
Hammers,  71.  • 

Hay,  43. 

Head-frame,  20-22. 
Historical  mention,  1. 


Hitches,  120. 

Hitch  timbers,  119,  122. 

Hoisting  apparatus,  20. 

engines,  20,  23,  24. 
Horden  sinking,  116. 
Hooks,  22. 
Hose,  78. 
Hydraulic  jacks,  85. 

Injector,  18. 

Kind,  97. 

Kind-Chaudron   boring  process,  97- 
102. 

Lackawanna  steel  sheet  piling,  39. 
Lagging,  36,  122. 
Lens  cementation,  103. 
Leyner  drill,  75. 
Lifting  water,  107-117. 
Lining  for  shafts,  118-139. 
Little  giant  drill,  73. 
Locking  through,  61. 

Mammoth  pump,  88. 
Man-lock,  60. 
Mixer,  concrete,  135. 

grout,  105. 
Moss  box,  98. 
Mt.  Lookout  freezing,  95. 

Non-rotating  rope,  25. 

Operation  of  concreting  shaft,   134- 

136. 

pneumatic  caisson,  62. 
sinking  shaft,  78. 

Packing,  122. 
Pile  driving,  40. 
Pile  hammer,  40. 
Piping,  17,  18,  110. 
Plant,  contractor's,  15. 

cost  of,  32. 

location  of,  32. 

for  sinking,  16-32. 
Pneumatic  process,  34,  59-63. 

limit  of,  60,  63. 


INDEX 


149 


Pneumatic  caisson,  59. 
Poetsch,  91. 
Poling  boards,  42. 
Powder  house,  30. 
Primary  power,  16. 
Progress,  sinking  caisson,  47. 

sinking  in  rock,  80,  81,  83. 
Prospect  holes,  9. 
Pulsometer,  112. 
Punch  blocks,  35,  121. 
Pumps,  sinking,  107,  113. 
Pumping,  107-117. 

Quadrilateral  shaft,  132,  134. 
Quicksand,  34,  60,  65. 

Rectangular  shaft,  2,  130,  132. 
Reheater,  27. 
Reenforcement,  44,  133. 
Relievers,  67. 

Rheinpreussen  colliery  sinking,  90. 
Riemer,  J.,  96. 
Ring  timbers,  121. 
Risk  of  water,  6. 
Rondout  caisson,  44. 
cementation,  105. 
Rope,  23,  25. 

Sack  borer,  86,  87. 

Safety  hook,  22. 

Saw,  29. 

Sealing  caisson  to  rock,  51,  63. 

Seals,  details  of,  52,  57. 

Secondary  power,  26. 

Sergeant  drill,  73. 

Sheet  piles,  37,  38. 

Shafts,  depths  of,  4. 

sizes  and  shapes  of,  2,  3. 
Shaft  bar,  75. 
Shaft  linings,  118-139. 
"Shaft  sinking  in  difficult  cases,"  96. 
Shield  method,  63-65. 
Shifts,  78. 

Shoes,  45,  48,  63,  85. 
Sinking-drum  process,  85-91. 
Sinking  pumps,  107,  113. 

in  rock,  66-84. 

in  surface,  33-65. 


Skin  to  skin  timber,  42. 
Slugger  drill,  73. 
Soft  ground,  36-65. 
Specifications,  13-15. 
Speed  of  sinking  caisson,  47. 

in  rock,  81,  83. 
Spoil,  5. 

Steam  hammer,  40. 
Steam  hose,  78. 
Steel  sheet  piles,  39. 
Straight-line  air  compressors,  26. 
"Straight  through"  air  lock,  62. 
Sump,  67. 
Supplies,  7,  31. 
Surface  ground,  33-65. 

Tamarack  shaft,  4. 
Thawing  dynamite,  30. 
Timber,  supply  of,  8. 
Timbering  firm  earth,  35. 

rock,  119-124. 

soft  ground,  36-42. 
Time  limit,  6. 
Tool  room,  31. 
Tools  for  sinking,  31. 
Tomson  system,  117. 
Tripods,  76. 
Tubbing,  English,  125. 

for  boring  process,  99. 

German,  128. 

Underground  water,  8. 
Underhanging  tubbing,  128. 

Ventilation,  79. 

Wages,  79. 
Wall  plate,  35,  121. 
Water  clause,  6,  11. 
Water  hoist,  4,  117. 
Water  in  surface,  33. 
Water  Leyner  drill,  75. 
Water,  pumping  of,  107-116. 
Water  rings,  15,  109. 

winding,  117. 
Wedges,  123,  127. 
Wedging  curb,  127. 
Wet  sinking  at  Conisboro,  112. 

Horden,  116. 


TN      Shurick  -  Coal  mine  surveying 


; 


MAY  I  9  19!  from  which  Itwasborrowed. 


Engineer!^* 
Library 


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